LIBRARY 

OF  THE 

UNIVERSITY  OF  CALIFORNIA. 

Class 


MINING  WITHOUT  TIMBER 


Published  by  the 

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New  York. 

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Publishers   of  Books  for 

Electrical  World  The  Engineering  and  Mining  Journal 

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Metallurgical  and  Chemical  Engineering  Power 


MINING 
WITHOUT  TIMBER 


BY 

ROBERT  BRUCE  ^RINSMADE,  B.  S.,  E.  M. 

FORMERLY  PROF.   OF  MINING  ENGINEERING  AT  WEST  VIRGINIA  UNIVERSITY,  MEMBER  OF  AMERICAN 

INSTITUTE   OF  MINING   ENGINEERS,   COAL  MINING  INSTITUTE  OF  AMERICA,  SOCIETY 

FOR  PROMOTION  OF  ENGINEERING  EDUCATION,  ETC. 


McGRAW-HILL    BOOK   COMPANY 

239  WEST  39TH  STREET,  NEW  YORK 

6  BOUVERIE  STREET,  LONDON,  E.  C. 

1911 


B 


COPYRIGHT,  1911 

BY 
McGRAw-HiLL  BOOK  COMPANY 


Printed  and  Electrotyfied 

by  The  Maple  Pres 

York.  Pa. 


Co 

MY     PARENTS 

THIS    BOOK   IS   GRATEFULLY 
DEDICATED. 


227177 


PREFACE 


The  rapid  depletion  of  the  primitive  forests  of  America  by  fires  and 
axe  in  recent  years  has  raised  the  price  of  wood  so  rapidly  that  the  mining 
industry  is  becoming  alarmed  as  to  its  future  supply  of  the  big  timber  of 
which  it  has  hitherto  been  such  a  prodigal  consumer.  Economy  in  the 
use  of  timber  has  been  essential  to  commercial  success  in  European 
mining  for  several  generations  and  it  was  to  the  Old  World  that  our 
operators  went  for  their  first  systems  of  timberless  mining. 

The  use  of  steel  and  masonry  for  the  support  of  mine  shafts  and 
tunnels  has  long  been  practised  in  Europe  not  only  because  of  dear  timber 
but  because  most  mines  there  are  considered  to  be  long-term  investments 
rather  than  temporary  speculations.  As  this  replacement  of  timber 
supports  by  other  material  involves  no  new  mining  system  and  has  been 
thoroughly  covered  by  other  writers,  it  will  not  be  described  in  this 
treatise.  Though  most  of  the  mining  methods  considered  consume 
some  timber,  their  economy  in  that  respect  is  so  marked  as  to  justify 
the  use  of  a  title  "Mining  without  Timber." 

The  work  is  not  intended  for  a  complete  treatise  on  mining,  but  is 
meant  to  deal  only  with  the  various  systems  of  excavation  with  such 
additional  matter  regarding  exploring,  blasting,  explosives,  and  the  con- 
trol of  ground  as  is  necessary  for  the  elucidation  of  the  main  theme. 
The  emphasis  is  placed  on  the  peculiar  problems  of  the  miner  and  little 
space  is  given  to  those  mining  topics  which  fall  chiefly  within  the 
provinces  of  the  mechanical,  constructing  or  electrical  engineer. 
Aqueous  excavation  by  hydraulicing,  by  dredging  and  by  solution  for 
such  deposits  as  those  of  placer-gold,  salt  and  sulphur  has  been  omitted 
because  that  subject  can  be  treated  best  in  special  treatises  of  which 
there  are  already  several  on  the  market.  A 

The  examples  of  practice  have  been  taken  mostly  from  North  Amer- 
ica, supplemented  by  a  few  from  Australia  and  South  Africa.  European 
practice  has  not  been  cited  not  only  because  its  valuable  features, 
modified  to  meet  American  conditions,  will  all  be  found  in  the  examples 
given,  but  because  the  subject  has  recently  been  specially  elaborated  in 
Mayer's  "  Mining  Methods  in  Europe."  Where  timbering  is  involved  in 
the  examples  its  details  have  been  condensed  since  framing  diagrams 
for  all  purposes  are  available  in  such  books  as  Storm's  "Timbering 
and  Mining."  The  costs  of  the  mining  are  mentioned  in  many  of  the 
examples  and  in  the  final  chapter  is  given  an  outline  of  the  manner  of 

ix 


X  PREFACE 

collecting  and  calculating  the  data  for  mine  evolution.  But  no  attempt 
has  been  made  to  treat  the  financial  side  of  mining  in  detail  for  that 
has  been  lately  comprehensively  done  in  modern  works  like  Ingalls' 
"Economics  of  Mining,"  Hoover's  " Principles  of  Mining"  and  Finlay's 
"  Cost  of  Mining." 

As  timberless  mining  systems  are  now  in  the  course  of  development, 
this  book  is  necessarily  somewhat  fragmentary  and  incomplete.  It  is 
merely  an  attempt  to  chronicle  the  generally  accepted  theories  and  the 
leading  examples  of  practice  so  that  the  student  may  learn  their  present 
status,  and  the  practising  engineer  may  have  access  to  the  record  of 
others'  experience  as  a  basis  for  the  solution  of  his  own  problems.  It 
aims  to  cover  mining  systems  broadly,  rather  than  particularly,  in  order 
to  be  equally  useful  to  both  coal  and  metal  miners. 

The  examples  of  practice  where  they  have  not  been  drawn  from  the 
author's  own  articles,  work  and  observations  in  the  mines  concerned, 
have  been  adapted,  as  acknowledged  in  the  Appendix,  from  articles 
published  recently  in  the  technical  press.  Thanks  are  due  the  editors 
of  "Engineering  and  Mining  Journal,"  "Mines  and. Minerals,"  "Mining 
and  Engineering  World,"  "Mines  and  Methods,"  "Mining  Science," 
"Mining  and  Scientific  Press,"  "Transactions  of  the  American  Institute 
of  Mining  Engineers"  etc.,  for  permission  to  republish  much  valuable  •' 
material  and  use  many  plates.  The  author  also  takes  this  opportunity 
to  express  his  gratitude  to  the  numerous  mining  men  including  mine- 
owners,  managers,  engineers,  accountants,  foremen,  and  miners,  whose 
unfailing  courtesy  to  him,  on  his  visits  of  investigation  to  their  mines, 
alone  has  made  this  book  possible. 

ROBERT  BRUCE  BRINSMADE. 
MORGANTOWN,  W.  VA. 
November,  1911. 


CONTENTS 

PAGE. 

PREFACE    .    ,    r  

CHAPTER  I. 

Explosives  and  Their  Use  in  Mining 1 

CHAPTER  II. 

Principles  of  Blasting  Ground 18 

CHAPTER  III. 

Compressed  Air  for  Mining 29 

CHAPTER  IV. 

Principles  for  Controlling  Excavations 33 

CHAPTER  V. 

Principles  of  Mine  Drainage 47 

CHAPTER  VI. 

Surface  Shoveling  in  Open  Cuts 62 

Example  1. — Moa  and  Mayari  Mines,  Cuba 62 

Example  2. — Mesabi  Range,  Minn 64 

Example  3. — Utah  Copper  Mine,  Bingham,  Utah 72 

Example  4. — Nevada  Con.  Mines,  Ely,  Nev 76 

Example  5. — Eastern  Pennsylvania  and  Illinois 82 

CHAPTER  VII. 

Surface  Mining      84 

Example  6. — Peurtocitos  Mine,  Cananea,  Mexico      84 

Example  7. — Mesabi  Iron  Range,  Minn 85 

Example  8. — Traders  Mine,  Iron  Mt.,  Mich 87 

Example  9. — Alaska  Treadwell  Mine,  Alaska 88 

CHAPTER  VIII. 

Underhand  Stoping 91 

Example  10. — Lead  Field  of  Southeast  Missouri 91 

Example  11. — Zinc-lead  Field  of  Southwest  Missouri 95 

Example  12. — Calumet  and  Arizona  Mines,  Bisbee,  Ariz 100 

Example  13. — Section  No.  21  Mine,  Marquette  Range,  Mich 104 

xi 


Xii  CONTENTS 

CHAPTER  IX. 

PAGE. 

Overhand  Sloping  with  Shrinkage.     No  Filling     .    .  106 

Example  14. — Wolverine  Mine,  Houghton  Co.,  Mich.  106 

Example  15.— Homestake  Mine,  Black  Hills,  So.  Dak.  108 

Example  16. — Gratz  Mine,  Owen  Co.,  Ky 112 

Example  17. — Alaska  Tread  well  Mine,  Alaska 113 

Example  18. — Veta  Grande  Mine,  Cananea,  Mexico  115 

CHAPTER  X. 

Overhand  Sloping  on  Waste  in  the  United  States 118-133 

Example  19. — South  Range  Mines,  Houghton  Co.,  Mich 118 

Example  20. — Minnesota  Mine,  Soudan,  Minn 121 

Example  21. — Superior  and  Boston  Mine,  Globe,  Ariz 124 

Example  22. — Metcalf  Mine,  Graham  Co.,  Ariz 128 

Example  23. — Copper  Queen  Mines,  Bisbee,  Ariz 130 

CHAPTER  XI. 

Overhand  Sloping  on  Waste  in  Mexico  and  Australia 134-154 

Example  24. — Los  Pilares  Mine,  Nacozari,  Mexico 134 

Example  25. — West  Australia * 143 

Example  26.— British  Mines,  Broken  Hill,  N.  S.  W 145 

Example  27.— Proprietary  Mine,  Broken  Hill,  N.  S.  W 152 

CHAPTER  XII. 

Overhand  Sloping  with  Shrinkage  and  Delayed  Filling 155-614 

Example  28.— Central  Mine,  Broken  Hill,  N.  S.  W 155 

Example  29. — King  Mine,  Graham  Co.,  Ariz 156 

Example  30. — Coronado  Mine,  Graham  Co.,  Ariz 158 

„     Example  31. — Los  Pilares  Mine,  Nacozari,  Mexico 162 

CHAPTER  XIII. 

Overhand  Sloping  with  Shrinkage  and  Simultaneous  Pillar  -caving    ....  165-180 

Example  32. — Miami  Mine,  Globe,  Ariz 165 

Example  33. — Boston  Con.  Mine,  Bingham,  Utah 171 

Example  34. — Duluth  Mine,  Cananea,  Mexico 176 

CHAPTER  XIV. 

Back-caving  into  Chutes  or  Chute-caving 181-191 

Example  35. — Hartford  Mine,  Marquette  Range,  Mich 181 

Example  36.— Pioneer  Mine,  Ely,  Minn 183 

Example  37. — Utah  Copper  Mine,  Bingham,  Utah 187 

CHAPTER  XV. 

Block-caving  System .  192-205 

Example  38. — Pewabic  Mine,  Menominee  Range,  Mich 192 

Example  39 — Mowery  Mine,  Santa  Cruz  Co.,  Ariz 194 

Example  40. — Detroit  Mine,  Morenci,  Ariz 196 

Example  41. — Commercial  Mine,  Bingham,  Utah 200 

Example  42. — Inspiration  Mine,  Globe,  Ariz 201 


CONTENTS  Xiii 

CHAPTER  XVI. 

PAGE. 

Slicing  Under  Mats  of  Timber  in  Barrels 206 

Example  43. — Old  Jordan  Mine,  Bingham,  Utah      ,    .  206 

Example  44. — Cumberland-Ely  Mine,  Ely,  Nevada      207 

Example  45. — Oversight  Mine,  Cananea,  Mexico 210 

CHAPTER  XVII. 

Slicing  Under  Ore  with  Back-caving  in  Rooms 214-227 

Example  46. — Goyebu,  Mesabi  and  Menominee  Ranges 214 

Example  47. — Mercur  Mine,  Mercur,  Utah 220 

Example  48. — Kimberley  Mine,  South  Africa 224 

CHAPTER  XVIII. 

Principles  of  Mining  Seams 228-237 

(a)  Comparison  of  Longwall  and  Pillar  Systems 228 

(b)  Comparison  of  Advancing  and  Retreating 229 

(c)  Mining  by  Roof-pressure 231 

CHAPTER  XIX. 

Advancing  Longwall  Systems  for  Seams 238-259 

Example  49. — Spring  Valley  Collieries,  111 238 

Example  50. — Montour  Iron  Mines,  Danville,  Pa 242 

Example  51.— Bull's  Head  Colliery,  Eastern,  Pa 245 

Example  52. — Vinton  Colliery,  Vintondale,  Pa 250 

Example  53. — Drummond  Colliery,  Westville,  N.  S .  255 

CHAPTER  XX. 

Pillar  Systems  for  Seams 260-275 

Example  54. — Advancing  System  Layouts 260 

Example  55. — Nelms'  Retreating  System 264 

Example  56. — Nelms'  Advancing-retreating  System 266 

Example  57. — Connellsville  District,  Western  Pennsylvania 268 

Example  58. — Pittsburg  District,  Western  Pennsylvania 272 

CHAPTER  XXI. 

Flushing  System  for  Filling  Seams  and  Recovering  Pillars 276-284 

Example  59. — Anthracite  District,  Eastern  Pennsylvania 276 

Example  60. — Robinson  Mine,  Transvaal 283 

CHAPTER  XXII. 

Comparison  of  Various  Mining  Systems 285-293 

CHAPTER  XXIII. 

Principles  of  Mine  Evaluation 294 

APPENDIX  I 301 

INDEX 303 


MINING  WITHOUT  TIMBER 


CHAPTER  I 
EXPLOSIVES  AND  THEIR  USE  IN  MINING 

An  explosion  may  be  defined  as  a  sudden  expansion  of  gas.  The 
substances  which  we  call  explosives  are  so  unstable  when  exposed  to  a 
suitable  flame  or  shock  that  they  suddenly  change  into  many  times 
their  original  volume  of  gas  with  the  evolution  of  heat.  If  the  change  to 
a  gas  takes  place  in  the  open,  there  is  a  flame  and  a  whiff  or  a  report.  It 
is  only,  however,  when  explosives  are  set  off  in  confined  spaces  like  drill- 
holes that  they  do  their  chief  work  in  mining.  Consequently  a  blast  or 
explosion  may  be  said  to  be  a  rapid  combustion  in  a  confined  space. 

Explosives  have  two  essential  constituents,  namely,  combustibles 
and  oxidizers.  They  may  be  broadly  divided  into  three  classes  accord- 
ing to  the  relation  which  the  combustibles  bear  to  the  oxidizers.  Class  I 
includes  the  mechanical  explosives,  or  thosa  in  which  the  ingredients 
constitute  a  mechanical  mixture;  class  II  includes  the  chemical  explo- 
sives or  those  in  which  the  ingredients  are  in  chemical  combination;  class 
III  inc  udes  the  mechanico-chemical  explosives  which  are  formed  of  a 
mixture  of  class  II  and  an  absorber. 

METHODS  OF  FIRING  EXPLOSIVES 

Explosives  are  set  off  by  two  means — ignition  and  detonation. 
Because  through  ignition  the  combustion  is  transmitted  by  heat  alone, 
it  gives  a  slower  explosion  than  one  started  by  detonation  which  trans- 
mits the  reaction  by  the  rapidity  of  vibrant  motion.  By  their  nature 
class  I  is  adapted  to  ignition,  and  classes  II  and  III  to  detonation. 

Ignition  is  commonly  performed  by  squibs,  fuse  or  electric  igniters. 
A  squib  is  really  a  self-impelling  slow  match,  made  by  filling  one-half  of  a 
thin  roll  of  paper  with  black  powder  and  the  other  half  with  sulphur. 
For  their  use  in  blasting,  a  drill-hole  ab,  Fig.  1,  is  loaded  with  an  explo- 
sive be  and  before  filling  the  hole  with  the  tamping  cd,  a  needle  ac  is 
inserted  into  the  explosive  so  that  when  it  is  withdrawn,  a  hole  of  a 
larger  diameter  than  the  squib  is  left  through  the  tamping  from  a  to  c. 


MINING    WITHOUT    TIMBER 


FIG.  1. — Drill-hole  section. 


The  squib  is  then  inserted  in  this  hole  with  the  sulphur  end  out,  and 
when  lit  the  slow-burning  sulphur  allows  time  for  the  miner  to  escape 
before  the  powder  of  the  squib  takes  fire  and  its  reaction  forces  the  squib 
along  the  holes  to  ignite  the  powder  at  c. 

A  fuse  is  merely  a  thread  of  black  powder  wrapped  with  one  or  more 
thicknesses  of  tape.     In  loading  the  hole,  Fig.   1,  the  fuse  would  be 

inserted  in  place  of  the  needle  ac.  A  fuse 
burns  commonly  at  the  rate  of  2  ft.  a 
minute.  Therefore  a  sufficient  length 
should  be  used  in  the  hole  to  allow  the 
miner  to  retire  in  safety,  after  splitting  and 
lighting  the  outer  end,  before  the  flame 
reaches  the  explosive  at  c. 

The  electric  igniter  consists  of  a  shell  a, 
Fig.  2,  enclosing  a  charge  of  fulminate 
mixture  in  b  and  of  sulphur  cement  in  e. 
The  copper  wires  c  pass  through  /  and 
enter  b  where  they  are  connected  by  a 
platinum  bridge  at  d.  For  ignition,  the 
shell  a  is  made  of  pasteboard  and  the 
igniter  is  placed  within  the  explosive  while 
the  wires  extend  outside  the  hole  to  a 
blasting  machine.  The  last  is  simply  a  small  armature  revolving 
between  its  poles  and  sending  a  current  through  the  igniters  in  the 
circuit  when  its  handle  is  shoved  down.  All  the  common  electric 
igniters  on  one  circuit  are  exploded  simultaneously,  but  a  recent  inven- 
tion is  a  delay-action  igniter  which  permits  electric  firing  in  sequence. 

Detonation  is  performed  by  fuse  and  cap  or  by  electric  caps.  A 
blasting  cap  is  simply  a  cylindrical  copper  cup  with  a  small  charge  of 
fulminate  mixture  in  its  bottom,  the  fuse  being  inserted  into  the  cup  and 
fastened  to  it  by  crimping  pincers.  The  cap  is  then  inserted  into  one 
cartridge  of  the  explosive  and  its  attached  fuse  tied  firmly  to  it  by  a 
string,  in  order  to  make  a  primer 
which  is  placed  near  or  on  the  top  of 
the  explosive.  The  loaded  hole  will 
then  resemble  Fig.  1,  the  explosive 
being  in  be,  the  cap  and  primer  at  c, 
and  the  fuse  along  ca.  Lighting  the 

fuse  is  the  same  as  for  ignition,  only  the  fuse  now  fires  the  cap  whose 
explosion  detonates  the  explosive. 

The  electric  cap  resembles  the  electric  igniter,  Fig.  2,  but  has  a  copper 
instead  of  a  pasteboard  case  a  and  the  quantity  of  charge  of  fulminate 
mixture  at  6  is  increased  as  the  sensitiveness  of  the  explosive  diminishes. 
The  electric  cap  is  inserted  in  and  fastened  to  a  primer-cartridge  like 


FIG.  2. — Electric  exploder. 


EXPLOSIVES    AND    THEIR    USE    IN    MINING  3 

fuse  and  cap,  the  electric  cap  being  fired  by  a  blasting  battery  in  the 
same  way  as  the  electric  igniter. 

LOADING  AND  TAMPING 

A  mechanical  explosive  like  black  powder  usually  comes  in  bulk. 
For  loading  it  is  poured  into  a  cartridge  (the  size  of  the  hole)  which  is 
made  by  rolling  a  piece  of  paper  around  a  pick  handle.  For  damp 
holes  the  cartridge  must  be  oiled  or  soaped  on  the  outside.  This  paper 
cartridge  is  pressed  down  into  the  hole  by  a  soft  iron  tamping  bar  whose 
tip  should  be  an  expanding  copper  cone  grooved  on  the  edge  for  the 
purpose  of  allowing  the  copper  loading  needle  or  fuse  to  pass.  Tamping 
bars  with  iron  tips  or  iron  needles  are  highly  dangerous  in  formations 
containing  pyrite  or  other  hard  minerals,  on  which  the  iron  might  strike 
a  spark,  and  their  use  is  therefore  prohibited  by  law  in  many  places. 

A  mining  explosive  of  class  II  or  III  is  handled  in  paper  cartridges 
which  can  be  ordered  of  a  diameter  to  fit  the  hole.  Before  loading  they 
are  slit  around  lengthwise  to  permit  of  the  explosive  taking  the  shape  of 
the  hole  when  it  is  pressed  down  by  a  tamping  bar  which  should  be  of 
wood  for  these  explosives,  instead  of  copper-tipped  iron,  on  account  of 
their  being  more  sensitive  to  any  shock  than  black  powder. 

In  coal  mines,  coal  dust  is  commonly  used  for  tamping  black  powder, 
but  this  is  a  very  unsafe  practice  in  dangerous  mines,  for  a  windy  or 
blown-out  shot  will  have  its  normal  flame  increased,  both  in  length  and 
duration,  by  the  ignition  of  the  tamping.  The  best  materials  for  tamp- 
ing are  a  fine  plastic  clay  or  loam  and  ground  brick  or  shale,  and  al- 
though sand  is  too  porous  to  do  well  for  black  powder,  it  answers  for 
higher  explosives  but  must  be  confined  in  paper  cartridges  for  use  in 
uppers. 

Water  is  used  as  tamping  for  nitro-glycerine  and  high  explosives  in 
wet  down-holes,  but  it  is  little  better  than  nothing.  The  fact  that  higher 
explosives  will  break  rock  without  any  tamping  has  caused  many  miners 
to  abandon  tamping  them  altogether  on  account  of  the  ease  of  recapping 
untamped  charges  in  case  of  a  misfire.  Mechanical  explosives  must  be 
tightly  tamped,  nearly  to  the  collar  of  the  hole,  or  they  will  blow  out 
instead  of  breaking  the  rock,  and  although  the  tamping  may  be  shortened 
with  detonating  explosives,  as  they  become  quicker  and  stronger,  a  short 
length  of  tamping  adds  to  the  efficiency  of  the  highest  explosives. 

Where  only  quick-acting  explosives  of  classes  II  or  III  are  at  hand 
and  it  is  desired  to  blast  with  the  slow  action  of  class  I,  the  object  can 
be  partially  obtained  by  special  methods  of  loading.  These  methods 
provide  an  air  cushion  between  the  explosive  and  the  rock  and  tamping 
by  either  having  the  stick  of  explosive  of  considerably  smaller  diameter 
than  the  drill  hole  or  by  having  a  very  porous  cellular  tamping  to  sepa- 
rate the  tight  tamping  from  the  explosive. 


4  MINING    WITHOUT    TIMBER 

Before  examining  the  various  mine  explosives  in  detail,  let  us  consider 
an  illustration  of  the  method  of  calculating,  from  the  chemical  equation 
of  an  explosive,  its  calorific  power,  its  temperature,  and  the  number  of 
expansions  and  its  consequent  exploding  pressure.  Let  us  assume  the 
simplest  case  of  a  mechanical  mixture  of  hydrogen  and  oxygen  at  -a 
temperature  of  0°  C.  and  at  sea-level  pressure  of  760  mm.  of  mercury. 
Then  the  chemical  equation  for  complete  combustion  is 

2H2  +  02  =  2H2O.  (1) 

the  molecular  weights  being  4  +  32  =  36.  (2) 

If  t  =  thermometer  temperature  in  degrees  centigrade  of  the  explosion; 

T  =  absolute  temperature  in  degrees  centigrade    of    the    explosion; 

2  =  sign  for  summation; 

WW^W2i  etc.  =  weights  in  grams  of  various  combustibles  of  the  explosive; 
00^2,  etc  =  calorific  power  in  calories  of  various  products  of  combustion 

of  the  explosive; 
wWiW2,  etc.  =  weights  in  grams  of  various  products  of  combustion  of  the 

explosive; 
ssvs2,  etc.  ==  specific  heat  in  calories  of  various  products  of  combustion  of 

the  explosive; 

V  —  volume  of  explosive  originally; 

Vl==  volume  of  explosive  due  to  chemical  reaction  alone; 
V2  =  volume  of  explosive  due  to  chemical  reaction  and  resulting  tempera- 

ture, t\ 

P  =  pressure  of  explosive  originally; 
P2  =  pressure  of  explosive  finally; 
then  we  have  from  thermo-chemistry, 

i    — 


etc. 

For  the  given  problem  we  have  from  equation  (2), 

W  =  4  grams  of  H  gas; 

w  =  3Q  grams  of  H.20  vapor. 
From  thermo-chemistry  we  have, 

C  =  28,780  cal.  for  H; 

s=  0.4805  cal.  for  H2O  vapor; 
substitute  in  (3)  and 


36X0.4805 

Then,  from  Avogardro's  law,  that  the  molecules  of  equal  volumes  of  all 
gases  under  like  conditions  occupy  the  same  volume,  we  have  from  (1), 

2  vols.  H  +  l  vol.  0  =  2  vols.  H20, 
or 

(4) 


EXPLOSIVES    AND    THEIR    USE    IN    MINING  5 

From  Charles'  law,  the  volumes  of  gases  vary  directly  as  their  abso- 
lute temperature  we  have  thus 


V,     0  +  273 
or 

-<   _6660F1 

}  2=  l7sT; 

substitute  from  (4)  and  we  have 

6J560X2F 
273X3 

From  Boyle's  law,  if  the  gas  of  volume  V2  is  prevented  from  expand- 
ing beyond  volume  V,  we  have  for  the  final  pressure  P2  in  the  explosive 
chamber  P, 

P       V 

/_2__    V  2 

P       V 

or 


Substitute  in  (6)  from  (5)  and,  as    P  =  l  atmosphere  =  14  .  7  Ibs.  per 
sq.  in.,  we  have 

16.2FP 
P2  =  ----  —  --------  =  16.2   atmospheres  . 

or  238  Ibs.  per  sq.  in. 

From  physics,  T==t  +  273, 
hence 

t  =  T—  273  =  6660—273  =  6387°  C. 

In  practice,  this  theoretical  pressure  and  temperature,  resulting  from 
the  explosion,  would  have  to  be  multiplied  by  a  fractional  factor  of 
efficiency  to  allow  for  imperfect  combustion  and  loss  of  heat  through 
radiation  and  leakage.  In  large  charges,  these  losses  are  proportionally 
less  than  in  the  case  of  small  charges.  This  fact,  coupled  with  the  greater 
likelihood  of  their  meeting  weak  places  in  the  blast's  burden,  accounts 
for  the  higher  efficiency  of  the  former.  These  theoretical  calculations 
are  especially  useful  in  comparing  the  relative  strength  of  different 
explosives  of  the  same  type.  In  France,  they  are  used  extensively  in  the 
inspection  of  permissible  explosives  to  determine  if  their  final  tempera- 
ture is  sufficiently  low  for  use  in  dangerous  coal  mines 

The  practical  usefulness  of  explosives  depends  upon  (1)  their  cost  of 
manufacture;  (2)  their  safety  and  convenience  as  regards  transportation 


O  MINING    WITHOUT    TIMBER 

and  storage;  (3)  method  necessary  for  their  loading  and  exploding; 
(4)  their  exploding  pressure;  (5)  the  rapidity  with  which  they  explode; 
(6)  the  length  and  temperature  of  the  flame.  These  six  factors  will  now 
be  discussed  seriatim.  Factor  (1),  or  the  cost,  is  often  the  most  impor- 
tant factor  in  commercial  operations  like  mining,  although  for  purposes 
of  war  it  is  often  little  considered.  Factor  (2)  or  safety,  affects  the 
desirability  for  all  purposes,  the  more  sensitive  the  explosive,  the  higher 
the  freight  rate  by  rail  or  boat,  and  if  sensitive  beyond  a  certain  point, 
it  cannot  be  shipped  thus  at  all.  Those  explosives  which,  like  dynamite, 
freeze  at  ordinary  winter  temperatures  are  at  a  disadvantage  as  are  also 
those  which,  like  black  powder,  are  handled  loose  and  can  be  easily 
ignited  by  a  spark  struck  by  a  hob-nailed  shoe  on  a  floor  spike.  Some 
explosives,  like  imperfectly  washed  guncotton,  are  liable  to  explode  by 
spontaneously  generated  heat,  while  others  become  dangerously  sensi- 
tive if  exposed  to  the  sun  during  shipment.  The  desirability  of  explo- 
sives belonging  to  either  of  these  last  two  mentioned  classes  is  plainly 
discounted  because  of  these  attributes.  The  next  factor  (3)  or  loading 
and  exploding,  is  important  in  connection  with  conditions  such  as  prevail 
in  dangerous  coal  mines  (where  an  open  light  is  prohibited),  in  subaque- 
ous blasting  (where  both  explosive  and  exploder  must  be  unaffected  by 
water),  or  where  misfires  could  not  be  corrected.  Factor  (4),  or  the 
pressure,  is  what  determined  the  real  effective  breaking  force  of  the 
explosion,  but  it  is  modified  in  practice  by  (5),  or  the  rapidity  of  the 
explosion.  Slow  and  fast  explosives  are  comparable  to  presses  and 
hammers  for  forging  steel.  The  former  exerts  its  pressure  gradually 
until  the  strain  exceeds  the  tensile  strength  of  the  material  and  the  rock 
gives  way  along  a  surface  of  fracture.  The  latter  gives  a  sharp  quick 
blow  which  will  shatter  the  surface  of  rock  exposed  to  the  explosive  before 
any  fracturing  action  is  exerted  on  the  blast's  burden  of  rock. 

The  slow  explosive  will  detach  the  rock  in  large  masses  while  the 
fast  type  may  crush  it  to  bits.  Black  powder  is  an  example  of  the  first 
and  nitro-glycerine  of  the  second.  Explosives  with  all  graduations  of 
rapidity  between  these  extremes  are  on  the  market.  The  fastest  explo- 
sives are  applicable  where  the  rock  is  very  hard  to  drill  as,  for  example, 
in  the  case  of  certain  Lake  Superior  hematites,  or  where  a  tremendous 
force  must  be  exerted  from  confined  spaces  as  in  breaking  the  cut  for 
development  passages;  also  where  a  shattering  rather  than  a  fracturing 
action  is  needed,  as  in  chambering  the  bottom  of  drill  holes  or  in  shooting 
oil  wells.  The  slowest  explosives  are  used  in  quarrying,  for  the  purpose 
of  detaching  monoliths,  or  in  consolidated  or  soft  rock  which  can  be 
fractured  by  a  slow,  pressing  movement  but  only  dented  by  a  quick 
hammer  blow. 

Factor  (6),  or  the  flame  and  temperature,  is  an  important  considera- 
tion for  blasting  in  gassy  or  dusty  coal  mines.  The  so-called  "permis- 


EXPLOSIVES    AND    THEIR    USE    IN    MINING  / 

sibles"  are  explosives  made  to  fall  below  a  minimum  legal  requirement 
as  regards  length  and  temperature  of  flame.  When  one  considers  that  a 
permissible  like  carbonite  gives,  in  practice,  a  flame  height  of  15.8  in. 
and  a  flame  duration  of  0.0003  seconds,  as  compared  with  50.2  in.  and 
0.1500  seconds  respectively,  for  black  powder,  we  can  see  how  much 
safer  the  permissible  is  to  use. 

We  will  now  consider  the  properties  of  the  three  classes  of  explosives : 

CLASS  I,  OR  MECHANICAL  EXPLOSIVES 

The  common  representatives  of  this  class  are  black  powder  and 
mechanical  permissible  explosives.  Black  powder  was  discovered  before 
600  A.  D.  by  the  Chinese,  and  by  Roger  Bacon  in  1270,  but  it  was  not 
used  for  mining  until  Martin  Weigel  introduced  it  at  Freiberg  in  1613. 
It  can  be  made  from  a  single  combustible,  charcoal,  mixed  with  an 
alkaline-nitrate  oxidizer,  but  in  order  to  lower  its  ignition  temperature 
for  blasting  to  about  275°  C.,  part  of  the  charcoal  is  replaced  by  sulphur. 
For  the  cheaper  blasting  powders,  the  oxidizer  is  sodium  nitrate  which, 
being  easily  affected  by  dampness,  is  replaced  in  the  higher  grade  powders 
by  potassium  nitrate.  The  ingredients  are  first  ground  then  mixed 
thoroughly  while  moist  and  finally  pressed  in  cakes,  dried,  broken  and 
sized.  Assuming  the  equation  for  the  complete  combustion  of  black 
powder  to  be. 

3C  +  S  +  2KN08  =  3C02  +  N  +  K2S.  (7) 

We  have  by  calculation  for  its  percentage  composition, 

carbon  =  13.4 

sulphur  =11.8 

sodium  nitrate  =  74. 8 

100.0 

and  for  the  percentage  composition  by  volume  of  its  resulting  gas, 

C0a  =  75 

N     =25 
100 

The  theoretical  exploding  temperature  is  4560°  C.  and  the  pressure  is 
5820  atmospheres.  In  practice  the  composition  is  varied  according  to 
the  experience  of  each  maker.  As  the  combustion  is  imperfect,  poison- 
ous and  combustible  gases  like  carbon  monoxide,  hydrogen  sulphide 
and  hydrogen  and  unpleasant  vapors,  like  the  sulphide,  sulphate,  hypo- 
sulphite, nitrate  and  carbonate  of  potassium,  are  given  off  by  the  explo- 
sion and  sometimes  render  breathing  or  the  carrying  of  open  lights  in 
the  fumes  a  dangerous  procedure.  In  fact,  Bunsen's  experiments  proved 


8  MINING    WITHOUT    TIMBER 

that  only  one-third  of  the  ignited  gunpowder  really  followed  the  reaction 
of  equation  (7). 

Black  powder  is  sold  in  grains  which  vary  in  size  from  the  fine  sporting- 
gunpowder  to  the  2-in.  balls  of  artillery  powder.  For  blasting,  the  grains 
vary  in  diameter  from  one-eighth  to  one-half  of  an  inch,  and  the  rapidity 
of  the  explosion  decreases  with  an  increased  diameter  of  grain.  The 
grains  should  be  of  uniform  size,  quite  dry  and  thoroughly  tamped  in 
the  hole  in  order  to  get  good  results.  The  specific  gravity  of  lightly 
shaken  black  powder  is  about  the  same  as  water.  Its  cheapness,  non- 
freezing,  comparative  safety  for  shipping  and  handling,  easy  explosion 
by  ignition  and  slow  action  are  the  favorable  qualities  of  black  powder 
which  cause  its  wide  use.  For  coal  mines  free  from  dangerous  gases  and 
dust,  it  is  a  better  explosive  than  detonating  permissibles  whose  quicker 
action  breaks  up  the  coal  and  injures  the  roof  more.  Black  powder  is 
rendered  inefficient  for  many  other  purposes,  however,  because  of  its 
necessitating  much  tamping,  its  low  power,  the  readiness  with  which  it 
is  spoiled  by  moisture  and  its  long  flame. 

Of  the  mechanical  permissibles  bobbinite  has  been  extensively  used 
in  England.  Its  percentage  composition  is, 

Potassium  nitrate  =  65 . 0 

Charcoal  =  20.0 

Sulphur  =  2.0 

Paraffin  wax=  2.5 

Starch  =   8.0 

Water  =   2.5 

100.0 

It  is  thus  chemically  very  close  to  black  powder  excepting  that  it 
contains  more  charcoal  and  less  sulphur  and  makes  up  that  discrepancy 
by  the  addition  of  wax,  starch  and  water.  The  lack  of  sulphur  raises  its 
ignition  temperature  while  the  wax  forms  a  waterproof  coating  for  the 
grains  of  powder.  The  starch  and  water  absorb  heat,  shorten  the  flame 
and  decrease  the  exploding  temperature  to  under  1500°  C.  It  is  handled 
in  compressed  cartridges  with  wax  coverings.  It  has  a  central  hole  to 
admit  the  fuse,  for  ignition  by  squib  is  not  allowed  in  dangerous  coal 
mines. 

CLASS  II,  on  CHEMICAL  EXPLOSIVES 

The  five  common  explosives  of  this  class  are  guncotton,  nitro-glycerine, 
nitro-gelatin,  fulminates  and  picrates.  They  all  contain  nitryl  (N02) 
and  their  detonation  is  made  possible  by  the  unstable  quality  of  nitryl 
compounds. 

Guncotton. — This  was  discovered  by  Schonbein  in  1846,  but  it  was 
little  used  until  it  was  found  that  its  dangerous  instability  was  not 


EXPLOSIVES    AND    THEIR    USE    IN    MINING  9 

inherent  but  due  solely  to  the  surplus  acid  left  in  its  tissue  by  imperfect 
washing  methods  during  its  manufacture.  The  equation  for  making 
it  is, 

C6H1005  +  3HN03  =  C6H705  (N02)  ,  +  3H20.  (8) 

cotton  +  nitric  acid  =  guncotton  +  water. 

The  ingredients  are  allowed  to  stand  in  a  cold  place  for  some  time 
before  the  washing  out  of  the  free  acid  is  begun. 
.    The  reaction  on  exploding  is, 

2C6H705(NO2)  3  =  3C02  +  9C02  +  3N2  +  7H20.  (9) 

Equation  (9)  shows  that  the  explosion  gives  no  solid  product  like  the 
K2S  of  equation  (7)  and  that  the  percentage  composition  by  volume  of 
the  resulting  gas  is, 

CO2  =  13.7 

CO  =  40.  8 

1ST  ==13.7 

H2O=31.8 


100.0 

By  the  method  of  calculation  already  explained,  it  is  found  that 
guncotton  theoretically  has  an  exploding  temperature  of  5340°  C.  and  a 
pressure  of  20,344  atmospheres. 

The  combustible  qualities  of  the  large  percentage  of  carbon  monoxide 
resulting  from  its  explosion  render  guncotton  unfit  for  use  in  coal  mines, 
and  its  poisonous  qualities  make  it  unsuitable  for  any  underground  use. 
For  surface  work,  it  is  very  powerful,  smokeless,  does  not  freeze  and  is  not 
volatilized  or  decomposed  by  atmospheric  temperature.  It  ignites 
between  270  and  400°  F.  and  if  unconfmed  will  then  burn  quietly.  When 
dry,  it  is  sensitive  to  percussion  and  friction,  but  under  water  it  is  insen- 
sible to  ordinary  shocks.  Immersed,  it  absorbs  from  10  to  15  per  cent. 
of  water,  but  even  then  it  can  be  exploded  without  drying  by  the  use  of 
an  extraordinarily  strong  detonator.  Its  chief  disadvantage  above 
ground  is  its  high  cost  and  the  fact  that  it  comes  in  hard  compressed 
cartridges  (specific  gravity  about  1.2)  which  fit  drill  holes  only  imper- 
fectly and  therefore  lose  in  efficiency.  For  any  destructive  work  without 
the  use  of  drill  holes,  like  demolishing  walls,  dams  and  the  like,  the  sharp, 
sledge-hammer  blow  of  its  explosion  renders  it  very  efficacious. 

Nitro-glycerine  or  "Oil."  —  This  was  discovered  by  Sabrero  in  1847, 
but  did  not  become  commercially  valuable  until  1863  under  the  direction 
of  Alfred  Nobel.  The  equation  for  its  making  is, 

C3H8O3  +  3HNO3  =  C3H503(NO2)  3  +  3H2O.  (10) 

glycerine  +  nitric  acid  =  nitro-glycerine  -f  water. 

Strong  sulphuric  acid  is  an  ingredient  of  the  mixture,  but  it  does  not 
take  part  in  the  reaction,  which  must  take  place  at  a  moderate  tempera- 


10  MINING    WITHOUT    TIMBER 

ture  to  be  safe.  The  resulting  "oil"  is  much  easier  to  wash  than  gun- 
cotton  and  consequently  is  cheaper.  It  is  a  yellow,  sweetish  liquid 
poisonous  both  to  the  blood  and  the  stomach.  Its  specific  gravity  is  1.6. 
Its  freezing-point  is  about  45°  F.  and  to  insure  against  freezing  the  tem- 
perature must  be  above  52°  F.  When  frozen,  it  is  insensible  to  ordinary 
shocks,  as  is  also  the  case  when  it  is  dissolved  in  alcohol  or  ether.  It  is, 
therefore,  commonly  shipped  either  in  tin  cans,  packed  in  ice,  or  in 
so  ution  in  wood  alcohol.  It  can  be  precipitated  from  the  latter  before 
use  by  an  excess  of  water. 

Nitro-glycerine  does  not  evolve  nitrous  fumes  until  230°  F.  As  it 
begins  to  vaporize  at  about  100°  F.,  it  is  important  in  thawing  it  not  to 
exceed  this  temperature.  Thawing,  therefore,  is  only  safely  done  by 
heating  the  explosive  over  a  water  bath  at  less  than  90°  F.,  or  by  leaving 
it  in  a  room  of  the  same  temperature  for  some  time.  The  explosive 
ignites  at  only  356°  F.  and  if  then  pure  and  free  from  all  pressure,  jar  or 
vibration,  it  will  burn  quietly.  These  safe-igniting  conditions,  however, 
are  difficult  to  obtain,  for  a  small  depth  of  liquid  causes  sufficient  pressure 
to  explode  it  when  ignited.  Thus  a  film  of  it,  heated  on  a  tin  plate, 
burned  without  an  explosion  only  if  under  one-fourth  inch  thick.  The 
exploding  temperature  is  380°  F.  This  24°  margin  above  the  igniting 
temperature  accounts  for  the  numerous  cases  of  conflagration  without 
explosion.  The  reaction  of  the  explosion  is, 

4C3H503(N02)3-12C02  +  02  +  3N2  +  10H20.  (11) 

From  equation  (11)  the  explosive  product  is  gaseous  and  its  percen- 
tage composition  by  volume  is 

CO2=  46.0 

O=     3.8 

N     =   11.8 

H20=  38.4 

100.0 

By  the  previous  calculating  method,  it  is  found  that  theoretically 
the  exploding  temperature  is  6730°  C.  and  the  pressure  is  29,107  atmos- 
pheres. From  the  fact  that  its  explosive  product  contains  no  carbon 
monoxide,  "oil"  can  be  used  underground,  but  only  when  mixed  with 
an  absorber.  .  Alone,  it  is  too  sensitive  to  be  safe,  while  being  liquid,  if 
unconfined,  it  would  leak  from  holes  in  porous  rock,  and  if  confined  in 
canisters  it  will  not  fill  the  drill  hole.  With  its  great  speed  and  strength 
it  also  tends  to  shatter  locally  any  enclosing  rock,  except  the  toughest, 
rather  than  detach  it.  These  characteristics  render  it  inefficient  for 
most  mining  work. 

For  shooting  oil  wells,  however,  its  shattering  quality  renders  it 
peculiarly  suitable.  For  this  purpose,  a  cylindrical  canister  of  a  diam- 
eter to  fit  the  well  and  containing  from  100  to  200  Ibs.  of  nitro-glycerine, 


EXPLOSIVES    AND    THEIR    USE    IN    MINING  11 

is  carried  to  the  well  swung  from  the  body  of  a  spring  buggy.  After 
filling  the  well  with  water,  the  canister  is  topped  with  a  cap  and  lowered 
to  the  proper  depths  by  a  rope,  along  which  a  weight,  called  a  "  go-devil," 
is  dropped  onto  the  cap  to  cause  the  explosion. 

Nitro-gelatin. — This  was  discovered  by  Nobel  in  1875  and  is  a  yellow- 
ish jelly  of  considerable  toughness,  but  easily  cut  with  a  knife.  It  is 
made  by  dissolving  guncotton  in  nitro-glycerine.  Authorities  differ  in 
the  proportion  of  guncotton,  some  recommending  only  7  per  cent.  To 
balance  all  the  free  oxygen  of  the  nitro-glycerine  by  the  excess  carbon 
of  the  guncotton  alone,  takes  87.3  per  cent,  of  the  former  to  12.7  per  cent, 
of  the  latter  and  gives  the  following  equation: 

9C3H503(N02)3  +  CttH7O2(N02)3  =  33C02  +  15N2  +  26H20.       (12) 

From  equation  (12)  the  percentage  composition  of  the  solely  gaseous 
product  is, 

C02  =  44.6 
N=  20.2 
H2O=35.2 

By  the  theoretical  calculation,  the  exploding  temperature  is  7080°  C. 
and  the  pressure  is  27,100  atmospheres.  The  last  figure  shows  nitro- 
gelatin  to  be  only  7  per  cent,  weaker  by  weight  than  nitro-glycerine, 
while  its  somewhat  higher  cost  is  due  to  its  guncotton  ingredient.  When 
used  alone  for  military  purposes,  about  4  per  cent,  of  camphor  is  dissolved 
in  the  nitro-glycerine  along  with  the  guncotton  to  make  a  product  called 
military  gelatin.  The  last  explosive  is  so  insensitive  that  it  can  be 
punctured  without  effect  by  a  rifle  bullet.  The  common  nitro-gelatin 
is  much  less  sensitive  than  No.  1  dynamite,  to  shock  or  friction,  and 
unaffected  by  a  short  immersion  in  water  at  158°  F.  and  by  an  8-day 
immersion  at  113°  F. 

It  will  not  exude  nitro-glycerine  under  a  high  pressure  or  any  atmos- 
pheric temperature.  Its  specific  gravity  is  1.6  and  it  can  be  set  off  only 
by  a  strong  detonation.  It  ignites  at  399°  F.  and  will  then  only  burn 
when  unconfined.  When  it  freezes,  which  is  between  35  and  40°  F.,  it 
becomes  more  sensitive  than  normally  owing  probably  to  the  partial 
freeing  of  the  nitro-glycerine  ingredient. 

Nitro-gelatin  is  now  used  for  mining  wherever  the  highest  power 
explosive  is  needed  and  is  especially  adapted  to  wet  or  subaqueous 
blasting,  either  alone  or  as  "gelatin"  dynamite. 

Fulminates. — Mercuric  fulminate  is  the  common  commercial  salt. 
It  is  made  as  follows  from  mercuric  nitrate  and  alcohol: 

Hg(N03)2  +  C2H30  =  Hg(CNO)2  +  3H20  +  20.  (13) 

The  explosive  reaction  is 

Hg(CNO)2  =  HgO  +  CO  +  C  +  2N.  (14) 


12  MINING    WITHOUT    TIMBER 

Equation  (14)  shows  that  mercuric  fulminate  is  a  poor  explosive 
because  it  produces  the  poisonous  fumes  of  HgO  and  CO  as  well  as 
unburned  carbon.  If  a  little  damp,  it  explodes  very  feebly  and  if  quite 
wet,,  not  at  all.  However,  its  non-freezing  quality,  its  quick  hammer- 
like  vibrant  explosion  and  its  uniform  sensitiveness  to  ignition  or  shock 
cause  its  use  as  the  chief  ingredient  of  percussion-cap  mixtures  for  deton- 
ating other  explosives.  Its  exploding  temperature  is  305°  F. 

Picrates. — These  salts  are  founded  on  picric  acid,  which  is  made  by 
mixing  carbolic  and  nitric  acid  according  to  the  equation, 

C6H60  +  3HN03-C6H3(N02)30  +  3H20.  ^  (15) 

Its  explosive  reaction  is 

CeH,(NO2)8O  =  H2O  +  H  +  6CO  +  3N.  (16) 

Picric  acid  comes  in  yellow  crystals  which  are  soluble  in  hot  water  or  al- 
cohol, and  melt  at  230°  F.  It  is  used  very  largely  in  dyeing.  It  is  expens- 
ive to  make  and  difficult  to  explode.  Equation  (16)  indicates  that  it 
produces  much  of  the  poisonous  carbon  monoxide  which  shows  incom- 
plete combustion  and  consequently  a  decreased  'power.  Picrates  are 
the  basis  of  the  military  explosive  lyddite,  but  the  recent  commercial 
failure  of  the  excellent  mining  picrate  joveite  may  discourage  future 
attempts  to  adapt  them  to  blasting. 

CLASS  III,  MECHANICO-CHEMICAL  EXPLOSIVES 

This  class  will  be  considered  under  five  groups:  (1)  guncotton;  (2) 
nitro-glycerine;  (3)  nitre-gelatin;  (4)  fulminate;  (5)  nitro-benzol.  Deto- 
nating permissibles  for  coal  mining  fall  mainly  under  groups  (2)  and  (5) 
and  will  be  considered  last. 

Guncotton  Group. — The  evaporating  of  guncotton,  after  it  has 
been  dissolved  in  a  suitable  solvent  such  as  alcohol  or  acetone,  produces 
a  hard,  horny  material  which  is  the  basis  of  most  modern  smokeless  gun- 
powder. Its  chief  blasting  powder,  however,  is  tonite  which  is  formed 
by  adding  enough  barium  nitrate  to  guncotton  to  just  completely  oxidize 
the  gases  caused  by  the  explosion  as  follows: 

4C6H705(N02)  3  +  9BaNO3  =  24C02  +  21N  +  14H20  +  9BaO.      (17) 

The  percentage  composition,  by  volume,  of  the  gaseous  product  of 
equation  (15)  is, 

CO2=   45.7 

N=   20.0 

H2O=  34.3 

100.6 

By  calculation,  the  exploding  temperature  is  3590°  C.  and  the  pressure 
is  10,300  atmospheres,  which  are  two-thirds  and  one-half,  respectively, 


EXPLOSIVES    AND    THEIR    USE    IN    MINING  13 

of  the  corresponding  figures  for  guncotton.  As  an  offset  to  lessened 
power  tonite  is  plastic,  cheaper  than  guncotton  and  50  per  cent,  denser. 
Its  harmless  fumes  adapt  it  to  underground  use  and,  like  dynamite,  it  is 
packed  in  paper  cartridges.  It  has  been  extensively  used  in  England, 
where  it  is  shipped  under  the  same  safety  regulations  as  black  bowder. 
It  is  hard  to  ignite  and  when  alight,  it  normally  burns  slowly  without 
explosion.  Tonite,  like  guncotton,  is  non-freezable  and  is  detonated 
only  by  a  strong  cap.  Potassium  nitrate  has  been  used,  instead  of 
bar  um  nitrate,  as  the  oxidizer,  in  another  guncotton  mixture  of  similar 
properties  which  is  called  potentite. 

Nitro-glycerine  Group. — These  mixtures  are  called  dynamites. 
They  were  introduced  by  Nobel  to  lessen  the  sensitiveness  of  nitro- 
glycerine and  at  the  same  time  retain  its  other  good  qualities.  The 
absorber  of  the  aoil"  is  called  the  "dope,"  which  may  be  selected  to  be 
either  inert  or  active  in  the  explosion. 

The  freezing  temperature  of  all  dynamite  is  that  of  nitro-glycerine,  as 
is  also  its  behavior  when  frozen  and  its  method  for  being  safety  thawed. 
Dynamite  that  does  not  leak  nitro-glycerine  under  the  conditions  under 
which  it  is  to  be  used  is  one  of  the  safest  explosives  known.  It  should 
not  be  shipped,  however,  in  rigid  metallic  cases,  which  accentuate  shocks 
and  vibrations,  but  in  wooden  boxes  in  paper  cartridges  packed  in  saw- 
dust. Thus  packed,  it  has  failed  to  explode  when  dropped  on  the  rocks 
from  a  considerable  height  or  when  struck  by  heavy  weights. 

Dynamite  can  be  heated  with  less  danger  than  nitro-glycerine.  If 
set  on  fire,  it  will  usually  burn  quietly  unless  unfavorable  conditions  are 
present.  If  the  dynamite  is  in  a  closed  box,  its  smoke  cannot  escape  and 
consequently  the  pressure  may  be  raised  enough  to  cause  an  explosion. 
If  caps  or  gunpowder  are  present,  the  fire  will  explode  them  and  the 
resultant  shock  will  detonate  the  dynamite,  If  the  heat  from  the  fire 
causes  the  "oil"  to  exude  from  the  cartilages,  this  "oil,"  if  under  a  static 
head,  will  explode  when  ignited,  as  explained  above.  Again,  the  heat 
from  the  burning  dynamite  may  heat  the  adjoiningunlighted  cartridges 
to  the  exploding  temperature  of  380°  F.  before  they  get  sufficiently 
exposed  to  the'  air  to  ignite.  Heated  gradually  in  the  open  so  much  of 
the  "  oil "  may  be  evaporated  that  a  mere  whiff  ensues  when  the  exploding 
temperature  is  finally  reached. 

In  spite  of  all  these  dangerous  contingencies,  several  instances  are  on 
record  where  several  tons  of  dynamite  have  burned  in  conflagrations 
without  exploding.  If  afire  in  cartridges,  it  burns  slowly  like  sulphur, 
but  if  loose  it  will  burn  quickly  like  chaff. 

The  dope  first  used  was  inert  infusorial  earth  or  kieselguhr,  which  will 
safely  absorb  three  times  its  weight  of  nitro-glycerine.  The  resulting 
kieselguhr  dynamite  when  strongest  contains  75  per  cent.  "  oil. "  It  is  a 
pasty,  plastic,  unctuous,  odorless  mass  of  a  yellowish  color  with  a  specific 


14  MINING    WITHOUT    TIMBER 

gravity  of  1.4.  The  effect  of  the  dope  is  to  cushion  the  " oil"  so  that  the 
shock  to  explode  it  must  be  stronger  as  the  percentage  of  dope  becomes 
greater.  It  is  not  possible  to  explode  kieselguhr  dynamites  which  con- 
tain under  40  per  cent,  of  "oil"  and  even  with  60  per  cent,  it  takes  a 
strong  cap. 

The  disadvantage  of  75  per  cent,  dynamite  is  the  exudation  of  "oil'' 
on  a  warm  day  or  under  water  so  that  dangers  may  arise  from  having  to 
deal  with  the  sensitive  "oil"  before  suspecting  its  presence.  It  is  thus 
ordinarily  unsafe  to  ship  or  use  and  the  60  per  cent,  strength  is  now 
commonly  sold  as  No.  1.  The  strength  of  kieselguhr  dynamite  is  almost 
equal  to  that  of  its  contained  "oil." 

The  active-dope  dynamites  have  no  such  narrow  limitations  as  the 
inert  types  and  not  only  may  numerous  absorbers  be  used,  but  the  per- 
centage of  nitro-glycerine  may  vary  from  4  to  70  per  cent.  These  explo- 
sives go  under  various  names.  The  common  active  absorbents  are  such 
combustibles  as  wood  meal  or  fiber,  rosin,  pitch,  sugar,  coal,  charcoal,  or 
sulphur,  and  such  oxidizers  as  the  alkaline  nitrates  or  chlorates.  The 
chemical  composition  of  the  oil-dope  mixture  should  be  such  as  to  give 
only  completely  oxidized  products  on  combustion.  The  strength  of  this 
type  is  equal  to  that  of  the  "oil"  plus  that  of  the  explosive  dope  when 
completely  burned.  In  other  words,  black  powder  mixed  with  enough 
"oil"  to  detonate  it  would  all  burn  as  shown  by  the  reaction  of  equation 
(7),  thus  giving  several  times  more  power  than  when  ignited  alone.  The 
density  and  appearance,  as  well  as  the  necessary  strength,  varies  with 
the  dope  and  the  percentage  of  "  oil. "  The  commercial  method  of  rating- 
dynamite,  by  its  percentage  of  "oil,"  is  misleading  as  no  account  is  taken 
of  the  varying  strength  of  the  explosive  dopes. 

Nitro-gelatin  Group. — A  mixture  of  .this  group  is  called  a  gelatin 
dynamite.  Somewhat  more  expensive  than  nitro-glycerine,  it  is  prefer- 
able wherever  the  highest  power  is  desired  and,  being  unaffected  by 
water,  it  is  the  best  powder  for  subaqueous  use.  It  is  more  plastic  and 
less  sensitive  than  common  dynamite  and  therefore  easier  to  load  and 
safer  to  transport,  but  it  requires  a  stronger  cap  for  exploding.  The 
military  powder  gelignite,  a  favorite  in  England  and  Japan,  and  forcite 
come  under  this  group. 

Fulminate  Group. — For  percussion-cap  filling,  mercuric  fulminate 
is  mixed  with  a  sufficient  amount  of  some  oxidizer  to  insure  complete 
combustion  on  exploding.  Alkaline-nitrate  oxidizers  may  be  used  but 
potassium  chlorate  is  the  favorite.  The  latter  gives  the  following 
exploding  reaction: 

Hg(CNO)2  +  KG1O3  =  HgO  +  KC1  +  2C02  +  2N.  (18) 

Equation  (18)  shows  that  potassium  chlorate  should  form  30  per 
cent,  by  weight  of  the  mixture,  which  also  contains  a  little  gum  to  give 


EXPLOSIVES    AND    THEIR    USE    IN    MINING  15 

coherence.     Caps  are  designated  by  numbers  or  letters  according  to  the 
amount  of  fulminate  contained.     The  common  series  is. 

Hg  (CNO)2 
Cap  No.  Grains. 

1  4.5 

2  6.0 

3  8.0 

4  10.0 

5  12.0 

6  15.0 
6.5  19.0 

7  23.0 

8  30.9 

The  larger  the  cap,  the  more  expensive,  but  if  the  cap  selected  is  too 
small  to  insure  perfect  detonation  of  the  explosive,  incomplete  combustion 
will  ensue  with  noxious  fumes  and  loss  of  power.  In  general,  dynamite 
requires  stronger  caps  as  the  percentage  of  nitro-glycerine  or  the  temper- 
ature decreases. 

Nitro-benzol  Group. — Although  nitro-benzol  contains  nitryl  it  does 
not  contain  sufficient  oxygen  to  be  an  explosive  and,  when  unmixed  with 
its  oxidizer,  it  can  be  shipped  as  an  ordinary  chemical.  On  this  account, 
the  nitro-benzol  or  Sprengel  group  is  especially  adapted  for  use  in  isolated 
places  far  from  dynamite  factories.  The  favorite  Sprengel  explosive 
is  rackarock,  which  is  a  mixture  of  nitro-benzol  with  the  chlorate  or 
nitrate  of  potassium  or  with  sodium  nitrate,  as  an  oxidizer.  By  mixing 
77.6  per  cent,  of  mononitro-benzol  with  22.4  per  cent,  of  sodium  nitrate, 
we  can  get  the  following  reaction  on  detonation : 

2C6H5(N02)  +  10NaN03  =  12C02  +  6N2  +  5H2O  +  5Na2O.         (19) 

From  equation  (19)  the  percentage  composition,  by  volume,  of  the 
gaseous  product  is, 

C02=  52.2 

N=   26.1 

H90=  21.7 


100.0 

By  calculation,  the  theoretical  temperature  is  5300°  C.  and  the 
pressure  is  13,800  atmospheres.  Unlike  ignited  black  powder,  rackarock, 
when  properly  detonated,  follows  closely  its  theoretical  reaction  which 
shows  harmless  gases  and  a  temperature  of  79  per  cent,  and  a  pressure  of 
47  per  cent,  of  the  figures  for  nitro-glycerine.  For  practical  use,  the 
oxidizer  of  rackarock  is  handled  alone  in  wax-paper  cartridges  and  the 
required  quantity  of  nitro-benzol  is  not  poured  into  a  cartridge  until 
just  before  charging  the  drill  hole. 


1(5  MINING    WITHOUT    TIMBER 

Detonating  Permissibles. — These  explosives  practically  all  contain 
either  nitro-glycerine,  nitro-gelatin,  nnitro-bezol  or  ammonium  nitrate 
as  the  detonated  ingredient  and  some  contain  two  or  more  of  them. 
Their  exact  composition  is  usually  kept  secret  by  the  manufacturers, 
but  they  must  pass  the  government  tests  for  temperature  and  flame: 
These  explosives  are  made  of  various  strengths  and  require  stronger  caps 
than  common  dynamites.  Detonation  means  a  quick  generation  of  a 
small  quantity  of  hot  gas  while  the  ignition  of  black  powder  means  the 
slow  production  of  a  large  quantity  of  impure  gases  and  vapors.  A  large 
quantity  of  fine,  unstable  salt  like  magnesium  carbonate,  of  a  steam- 
generating  salt  like  ammonium  nitrate,  or  of  a  substance  with  much 
hygroscopic  moisture  like  wood  meal,  are  the  ingredients  relied  upon  to 
cool  the  quick  small  flame  of  permissibles.  The  compositions  of  a  few 
typical  permissibles  are  as  follows: 

Name.                               Nitro-benzol.  NH<N(>2             Ground  Wood.        Water. 

Amvis 4.50  90.0  5.0  0.50 

Ammonite r 12.00  88.0 

Electronite 19.0  (BaNo),. .  73.0  7.5  0.50 

Westfalit,    No.    1.    4.5  (rosin)3.. .  95.0  0.50 

Bellite,  No.  3 5.25  94.0  ...  0.75 

Carbonite 25.0("  oil  ")  .  .  O.SO(soda)  34.0  (NaNOs)  40.5 

MISFIRES 

The  cause  of  misfire  depends  upon  both  explosive  and  the  manner 
of  firing.  The  three  classes  of  explosives  with  their  methods  of  firing 
will  now  be  considered. 

Mechanical  Powders  of  Class  I. — In  breaking  coal  with  igniting  pow- 
ders, it  is  inadvisable  to  attempt  to  use  a  missed  hole  if  the  tamping 
must  first  be  dug  out,  therefore  a  new  hole  is  bored,  charged  and  fired 
alongside  the  first.  In  rock  breaking,  where  boring  holes  is  expensive, 
the  tamping  may  be  dug  out  safely  if  only  copper  tools  are  used  when 
approaching  the  powder.  However,  if  the  explosives  are  well  selected, 
and  kept  dry,  and  care  is  taken  in  locating  and  loading  the  holes,  misfires 
will  seldom  occur. 

With  squib-ignition  misfires  may  be  caused  by  (a)  wetness  of  powder; 
(6)  dampness  of  squib;  (c)  loss  of  powder  from  squib;  (d)  squib-hole 
clogged  by  dirt;  (e)  hole  too  long  for  squib  to  recoil  and  reach  powder. 

With  fuse-ignition  misfires  may  be  due  to  (a)  damp  powder;  (6)  cut- 
ting of  fuse  in  tamping;  (c)  imperfect  fuse;  (d)  damp  fuse;  (e)  loss  of 
powder  from  end  of  fuse. 

With  ignition  by  electric  igniter  misfires  may  occur  from  (a)  imperfect 
igniter;  (6)  damp  igniter;  (c)  wire  broken  in  tamping;  (d)  circuit  im- 
perfectly wired;  (e)  current  leakage  from  poor  insulation;  (/)  current 
deficiency  from  imperfect  or  overloaded  blasting  machine.  The  com- 


EXPLOSIVES    AND    THEIR    USE    IN    MINING  17 

pleteness  of  the  circuit  can  be  tested  before  the  exploding  by  passing  a 
feeble  current  through  a  galvanometer  placed  in  the  circuit. 

Detonating  Powders  of  Classes  II  and  III. — In  breaking  coal  with 
these  powders  it  is  better,  as  with  igniting  powders,  to  bore  and  load  a 
new  hole  than  to  dig  out  the  tamping  from  a  missed  hole.  In  rock  work, 
it  is  good  practice  to  dig  out  the  tamping  from  a  missed  hole  to  within 
only  half  an  inch  of  the  powder  and  then  insert  a  new  primer  cartridge 
with  detonator  and  retamp.  The  excavation  of  tamping  should  be 
cautiously  done  when  approaching  the  powder  and  care  be  taken  not 
to  strike  the  cap. 

Dynamite  should  not  be  allowed  to  remain  long  before  firing  in  water 
holes,  for  ths  water  may  displace  the  "oil"  and  perhaps  cause  a  misfire 
or  the  escape  of  "oil"  into  adjoining  crevices  when  it  may  later  be 
struck  by  a  pick  or  drill  and  explode.  Powder  should  never  be  used 
when  even  partly  frozen,  for  the  thawed  portion  may  explode  alone  and 
leave  the  frozen  residue  in  the  hole  or  blow  it  out  into  the  muck  to  become 
in  either  case  a  source  of  danger  for  the  next  shift  of  miners. 

In  firing  a  round  of  holes  in  sequence,  the  explosion  of  one  hole  may 
blow  off  the  primer  of  an  adjoining  hole  whose  remaining  charge  is  there- 
fore left  unexploded  in  the  hole-stump.  Except  for  the  last  contingency, 
and  that, of  two  holes  exploding  simultaneously,  the  counting  of  the  ex- 
ploding reports  gives  a  check  on  detonating  in  sequence  which  is  lacking 
in  simultaneously  firing  by  electricity.  An  electric  cap  may  be  damp  and 
conduct  the  current  through  the  circuit,  without  exploding  itse'f,  and  a 
missed  hole  will  thus  result.  A  fuse  may  have  a  broken  thread  of  powder 
whose  wrapping  may  catch  fire  and  smoulder  some  time  before  igniting 
the  powder  beyond  the  break  For  all  these  reasons  the  stumps  of 
blasted  holes  shou'd  be  carefully  examined  before  resuming  work,  and 
where  misfires  are  suspected  a  half-hour  interval  should  elapse  before 
revising  the  broken  face. 

FUSQ  and  cap  detonation  has  the  last  four  causes  of  misfires  already 
given  for  fuse  ignition  and,  in  addition,  is  liable  to  failure  of  the  cap, 
either  from  dampness,  imperfection,  or  insufficient  strength  for  the  given 
explosive. 

The  causes  of  misfires  already  given  for  electric  ignition  can  be  made 
to  read  correctly  as  the  causes  with  electric  detonation  by  simply  substi- 
tuting the  word  cap  for  igniter  and  adding  the  requirement  that  the  cap 
must  be  of  adequate  strength. 


CHAPTER  II 
PRINCIPLES  OF  BLASTING  GROUND 

It  is  only  in  recent  years  that  engineers  have  had  much  to  do  with 
the  details  of  underground  excavation,  as  it  was  thought  that  all  the 
schooling  necessary  for  the  successful  miner  could  be  gained  by  practice 
with  a  drill  and  shovel.  It  is  evident,  however,  that  where  rock  breaking 
forms  such  an  important  item  of  expense  as  it  does  in  most  mines,  it 
will  well  repay  study  to  ascertain  if  science  cannot  duplicate  here  the 
same  success  it  has  gained  over  empiricism  in  other  departments. 

After  an  explosion  of  powder  in  the  bore  hole,  Fig.  3,  the  sudden  ex- 
pansion of  the  resulting  gases  will  exert  its  force  equally  in  all  directions 
on  the  bore  hole,  until  either  the  enclosing  rock  or  the  tamping  yields 


/      \ 

\ 

/                1 

\ 

/                 -) 
1                      1 

\ 

r"^ 

A 

\ 

/  | 

\ 

'        /     ' 

--a   (- 

^' 

t' 

^N  ! 

/                    \  | 

\  i 

/                       1  1 

\i 

'                        1  1 

\i 

~-4 

V 

?       , 
/ 

Sectioa  Plan 

FIG.  3. — Cones  of  blasting  rock. 

and  the  gases  escape.  The  rock  will  yield  along  what  is  called  the  line 
of  least  resistance,  which  would  be  be  in  the  assumed  homogenous  rock 
of  Fig.  3.  It  is  evident  that  the  angle  0,  which  the  hole  ab  makes  with 
the  exposed  surface  or  the  free  face  of  the  rock,  can  vary  from  nothing 
to  90  deg.  At  0  =  0  deg.,  there  would  be  no  hole  and  at  90  deg.  the  hole 
would  be  in  the  position  be,  the  line  of  least  resistance,  and  would  give  a 
blown-out  shot.  The  quantity  of  rock  thrown  out  by  the  explosion  would 
have  the  volume  of  a  cone  with  an  altidue  be  or  h,  and  a  base  with  a  radius 
ac,  whose  volume  v  =  l/3  hn(ac)2  and  where  0  =  45  deg.  (the  usual  con- 
dition for  the  maximum  volume)  ac  =  h  and  we  have  v  — =  (nearly)  h3, 


or  if  m  is  a  constant,  depending  on  rock,  then  v  =  mh3. 

18 


PRINCIPLES  OF  BLASTING  GROUND  19 

For  a  case  with  two  free  rock  faces  if  the  powder  charge  be  placed  at 
e,  with  the  lines  of  least  resistance  eg  and  em  of  equal  length,  the  ex- 
plosion will  break  out  two  cones  def  and  fek,  or  nearly  double  the  volume 
for  one  free  face,  so  that  v  =  2mh3.  It  is  similar  for  three  or  more  free 
faces,  so  that  as  a  general  equation  we  have,  if  n  =  the  number  of  free 
faces,  v  =  nmh3. 

From  this  formula  it  can  be  seen  that  a  system  of  mining  should 
be  adopted  which  utilizes  as  many  free  faces  as  possible  in  breaking. 
In  development  work  for  vertical,  horizontal  or  inclined  drives  or  pas- 
sages, we  start  each  round  of  holes  with  one  free  face  and  With  our  cut 
holes  break  out  either  a  cone  or  a  wedge  whose  surface  forms  another 
free  face  for  the  benefit  of  the  other  holes  of  the  round.  In  stoping 
work,  which  must  be  started  from  a  drive,  we  can  always  manage  to 
maintain  two  and  often  three  free  faces  in  homogeneous  rock,  and  in 
stratified  formations  sometimes  four  or  more  faces,  as  a  bedding  plane 
is  often  nearly  the  equivalent  of  a  free  face. 

In  stratified  formations  the  correct  principles  of  breaking  are  especi- 
ally important  for  economy's  sake.  The  simpliest  case  is  that  of  beds 
2  to  4  ft.  thick.  Here  the  holes  should  be  drilled  in  a  plane  parallel 
to  the  beds  because  it  is  evident  that  we  can  more  easily  separate  two 
wet  coins  on  a  table  by  sliding  one  sideways  than  by  trying  to  lift  it 
off  vertically.  'Also  these  parallel  holes  do  not  weaken  the  blast  by 
allowing  the  powder  gases  to  escape  through  the  bedding  seam.  Where 
the  beds  are  thin,  say  under  8  in.,  we  encounter  the  possibility,  with 
holes  parallel  to  the  bedding,  of  having  only  the  small  bed  blown  out  that 
contains  the  hole.  For  this  reason  it  is  advisable  to  first  make  a  cut 
by  driving  holes  across  the  bedding  planes  and  then  break  to  the  cut 
with  the  balance  of  the  holes  drilled  parallel  to  the  bedding  plane,  but 
which  now  exert  their  maximum  force  perpendicular  instead  of  parallel 
to  the  beds. 

The  method  of  firing  also  affects  the  pointing  and  the  necessary 
number  of  holes  to  drill  for  breaking.  There  is  a  great  advantage  in 
simultaneous  or  electric  firing  wherever  a  weak  roof  or  the  greater 
danger  from  misfires  w  th  unskilled  miners  do  not  militate  against  it. 
In  Fig.  3  it  is  evident  that  only  the  cone  abd  and  the  double  cone  dek 
would  be  broken  out  by  the  charges  at  b  and  e  fired  separately,  but  if 
b  and  e  are  not  too  far  apart  and  are  fired  together  the  line  of  detach- 
ment will  be  along  the  lines  abek  instead  of  abdek  and  the  extra  volume  bde 
will  be  broken  with  no  extra  powder  or  drilling.  In  any  case  of  breaking, 
the  pressure  p  produced  by  the  explosive  multiplied  by  the  area  of  its 
section  a  (taken  along  the  axis  of  the  hole)  must  equal  the  ultimate 
tensile  or  shearing  strength  T  of  the  rock  multiplied  by  the  area  of  its 
surface  of  fracture  S  or  pa  =  TS. 

If  pa  is  greater  than  TS  it  means  an  excess  of  explosive  over  that 


20 


MINING    WITHOUT    TIMBER 


required  for  detaching  the  burden.  This  excess  causes  a  " windy" 
shot,  resulting  in  a  greater  air  blast,  a  louder  report  and  a  longer,  hotter 
flame  than  from  a  normal  shot.  A  normal  charge  leaves  traces  of  the 
drill  hole,  but  an  insufficient  charge  leaves  "candlesticks"  in  the  rock 
and  loose  pieces  of  the  burden  have  to  be  blasted  off. 


Plan  for  (a)  (b)  &  (c) 


UNDERGROUND  DEVELOPMENT 

In  illustrating  we  will  take  the  case  of  driving  horizontal  headings  or 
drifts  as  the  same  principles  of  breaking  apply  equally  well  for  inclined 
and  vertical  shafts  and  raises.  The  practical  difference  in  the  latter 
arises  from  the  setting  of  the  drills  and  the  handling  of  the  muck  and 
the  water,  and  the  fact  that  the  length  of  the  section 
in  shafts  generally  makes  the  central  cut  advisable. 
We  will  also  assume,  to  simplify  the  illustrations,  a 
heading  small  and  soft  enough  to  allow  its  breakage 
^y  rounds  of  nine  holes  in  three  rows  of  three  holes 
each,  although  often  nine  holes  are  more  effective  in 
four  rows,  one  of  three  and  the  balance  of  two  holes 
each.  For  longer  headings  with  harder  rock,  the  same 
principles  would  apply,  but  more  holes  must  he  added 
for  breaking  the  round.  On  this  basis  we  will  now 
consider  the  following  six  cases  of  formation. 

Case  I.  Homogeneous  Rock  Free  from  Bedding 
Planes  or  Joints  in  the  Face  of  the  Heading.  —  Since 
this  formation  breaks  equally  well  in  any  direction, 
the  holes  should  be  placed  for  the  most  convenient 
drilling  and  mucking.  For  setting  the  bar  horizont- 
ally, as  is  usual  where  it  is  desired  to  begin  drilling 
before  the  muck  from  the  last  round  is  cleaned  up, 
Sec.  (c)  the  placing  of  Fig.  4  (a)  is  a  favorite.  Here  the  ad- 

FIG.  4.—  Holes  for  head-  justable  arm  is  unnecessary  and  the  first  setting  of 

ings  with  horizontal  bar.       ,        .  ,    .„    ,  ,  ,         .,  „ 

the  bar  is  at  A  to  drill  the  upper  and  middle  rows 
with  the  machine  above  the  bar.  The  second  setting  of  the  bar  is  at 
B  and  the  machine  is  turned  under  it  for  drilling  the  bottom  row  3 
of  lifters.  The  horizontal  rows  of  holes  are  usually  fired  in  the  order 
1,  2,  3,  Fig.  4  (a). 

The  side  instead  of  the  bottom  cut  is  handiest  if  we  wish  to  set  the 
bar  vertically.  We  first  set  up  at  A,  Fig.  5,  and  drill  row  1,  then  at  B 
with  the  machine  on  one  side  to  drill  row  2  and  on  the  other  to  drill  row 
3.  Here  the  vertical  rows  of  holes  are  fired  in  the  order  1,  2,  3,  Fig. 
5  plan.  In  other  to  keep  the  passage  straight,  the  cut  holes  of  row  1 
will  be  put  for  the  next  round  on  the  opposite  side  to  what  is  shown, 
so  that  the  finished  sides  have  a  zig-zag  appearance,  alternately  right  and 


See.  (b) 


PRINCIPLES    OF    BLASTING    GROUND 


21 


left  as  shown  in  the  plan  of  Fig.  5.  The  middle  hole  of  vertical  row  1 
points  downward,  like  hole  c,  instead  of  flat-wise,  like  the  balance  of 
horizontal  row  1,  so  as  to  throw  out  a  bottom  cut  and  avoid  a  horizontal 
inclination  to  the  face  too  acute  for  rapid  progress  in  a  narrow  heading. 
For  large  tunnel  headings,  8  ft.  square,  in  hard  homogeneous  rock,  the 
cone  or  "Leyner"  center-cut  system  has  recently  permitted  of  very 


Plan 


2 

e 

i 

S3 
^^jL 

Section 
FIG.  5. — Holes  for  headings  with  vertical  bar. 

fast  driving  in  western  metal  mines.  It  is  especially  adapted  to  the 
water  Leyner  drill  on  account  of  the  many  upper  holes  used  and  the 
fact  that  th's  drill  is  short  enough  to  allow  the  sharp  pointing  of  the 
holes  with  two  settings  of  the  bar.  For  hard  steel  ore  and  jasper  in  a 
Michigan  iron  mine,  this  system  was  thus  applied. 

In  Fig.  6,  A  is  the  bar  in  first  position  for  two  machines  and  from  its 


K 8-0- H 


10b      12* 


16r 


Front  Elev. 

FIG.  6. — Holes  for  Leyner  tunnel  cut. 

top  the  four  back  holes,  Nos.  9,  10,  11  and  12,  are  drilled.  The  machines 
are  then  tipped  forward  until  the  crank  can  just  turn  and  clear  the  back 
or  top  of  the  drift  for  drilling  the  top  center  cut  holes  Nos.  1  and  2, 
while  finally  they  are  turned  under  the  bar  for  side  holes  Nos.  5,  6,  7 
and  8.  The  bar  is  then  changed  to  position  B,  the  machines  are  set  up  on 
top  and  side  holes  Nos.  13  and  14  are  drilled.  Then,  after  turning  the 


22  MINING    WITHOUT  TIMBER 

machines  under  the  bar,  they  are  tipped  up  in  front  so  the  crank  just 
clears  the  bottom  of  the  drift  and  holes  Nos.  3  and  4  are  drilled  about  to 
meet  Nos.  1  and  2  in  the  center  of  the  heading.  The  four  lifters,  Nos. 
15,  16,  17  and  18,  are  the  final  holes.  In  softer  and  better-breaking 
ground,  cut  holes  Nos.  5  and  6,  one  lifter  and  one(back  hole  can  be  left 
out,  but  the  four  cut-holes,  IJos.  1,  2,  3  and  4,  are  nearly  always  used 
and  are  pitched  up  and  down  and  in,  to  meet  about  in  the  center. 

The  five  remaining  cases  are  given  for  regularly  stratified  rock,  but 
the  joints  or  cracks  of  massive  rock  may,  like  bedding  planes,  often  be 
utilized  for  breaking. 

Case  II.  Rock  in  Horizontal  Beds;  (a)  Medium  Thick  Beds. — Here 
the  best  results  from  the  powder  can  be  got  by  two  settings  of  the  bar 
vertically  and  following  the  drilling  and  firing  directions  given  above 
for  the  method  illustrated  by  Fig.  5.  In  the  disseminated  lead  mines 
of  southeastern  Missouri  (Example  9,  Chapter  VIII),  this  method  is 
modified  as  follows: 

For  a  drift  10  ft.  wide  by  6  1/2  to  7  ft.  high,  12  to  13  holes  are  needed, 
placed  in  three  rows  horizontally  by  four  rows  vertically.  The  bar  is  set 
up  once  to  drill  each  vertical  row  of  holes,  four  set-ups  being  necessary 
to  complete  a  round.  Each  vertical  row  is  fired  separately  by  fuse  and 
dynamite  and  as  only  three  or  four  holes  are  fired  at  a  time,  not  enough 
smoke  or  broken  rock  is  produced  to  prevent  the  drillers  from  setting 
up  again  very  soon  after  blasting.  This  method  with  three  shifts  of  two 
drill  men  each  allows  an  advance  of  5  to  7  ft.  in  24  hours  with  2  3/4-in. 
drills.  By  the  former  center-cut  system,  two  drills  and  four  men  were 
able  to  advance  only  10  to  15  per  cent,  faster  than  by  the  one  drill  and 
the  side-3ut  method  just  described,  all  loading  and  tramming,  in  each 
case,  having  been  done  by  muckers.  . 

(6)  Thin  Beds. — Here,  as  already  explained,  the  cut-holes  must  cross 
the  bedding  planes.  A  bottom  cut  is  advisable.  The  bar  is  set  horizon- 
tally at  A,  Fig.  4  (6).  Often  all  three  rows  can  be  drilled  direct  although 
sometimes  the  use  of  the  adjustable  arm  on  the  bar  is  necessary  to  get 
the  correct  pointing  of  the  holes.  The  holes  of  row  1  are  fired  first  and 
break  out  the  cut  to  the  bedding  plane  on  the  floor  of  the  heading.  Be- 
fore loading  the  row  of  cut-holes,  it  is  often  helpful  to  stop  up  their  bed- 
ding planes,  around  the  powder,  with  clay  but  this  precaution  is  unneces- 
sary in  the  two  upper  rows  where  the  holes  are  parallel  to  the  beds. 

Case  III. — Rocks  in  Vertical  Beds  Parallel  to  Heading;  (a)  Medium 
Thick  Beds. — This  case  requires  the  bottom  cut  of  Fig.  4  (a)  which  has 
already  been  described  under  Case  I.  The  use  of  this  method  in  the 
vertical  copper  veins  of  Butte,  Mont.,  is  as  follows:  The  placing  of  holes 
is  shown  in  Fig.  7  for  the  12-hole  system,  although  for  most  rock  nine 
holes  are  ample,  the  center  holes  of  rows  2,  3,  and  4  being  omitted.  For 
this  arrangement  the  drill  bar  (with  adjustable  arm)  need  only  be  set 


PRINCIPLES    OF   BLASTING    GROUND 


23 


up  once  vertically,  as  shown.  The  round  of  holes  is  usually  loaded  and 
fired  at  one  time  and  goes  off  in  the  order  of  1,  2,  3,  4.  Some  of  the 
miners  regulate  the  explosions  by  cutting  the  fuse  of  different  lengths 
and  spitting  them  simultaneously  while  held  together  in  the  hand,  and 
others  by  cutting  all  the  fuse  of  the  same  length  and  spitting  them 
separately  in  the  required  order. 

(6)  Thin  Beds. — The  solution  of  this  case  follows  Fig.  5  and  also 
resembles  Case  II  (b)  except  that  here  the  side  instead  of  the  bottom  cut 
is  used.  With  one  setting  of  the  bar,  the  three  vertical  rows  K,  2  and  3 
may  be  drilled  and  shot  in  the  same  order,  row  K  breaking  out  the  cut, 
along  a  side  bedding  plane,  mn,  and  rows  2  and  3  breaking  to  the  cut. 


Long-it.  Section 

FIG.  7. — Holes  for  sloping. 


Here  it  is  not  so  necessary  for  alignment,  as  in  Case  II  (a),  to  alternate 
the  cut  on  each  side  of  the  heading,  but  it  is  often  an  advantage  especially 
where  the  vertical  bedding  planes  are  ill  defined. 

Case  IV. — Rocks  in  Vertical  Beds  Cutting  the  Heading  at  an  Angle;  (a) 
Medium  Thick  Beds. — If  the  cutting  angle  which  the  bedding  plane 
makes  with  the  side  of  the  heading  is  45  deg.  or  less,  the  method  of 
Fig.  4  (a)  is  usually  preferable.  If  the  cutting  angle  is  more  than  45 
deg.,  the  choice  between  the  methods  of  Fig.  4  (a)  and  of  Fig.  5  will 
often  be  merely  a  question  of  convenience  in  setting  the  bar  horizontally 
or  vertically,  respectively. 

(b)  Thin  Beds. — With  a  cutting  angle  of  45  deg.  or  less  the  method 
of  Fig.  5  is  the  best.  Where  the  cutting  angle  is  more  than  45  deg., 
the  choice  between  the  methods  of  Fig.  4  (a)  and  Fig.  5  depends  on  set- 
ting the  bar  as  in  Case  IV  (a). 


24  MINING    WITHOUT    TIMBER 

Case  V. — Rocks  in  Inclined  Beds  Dipping  Toward  the  Floor  of  the 
Heading. — For  either  medium  thick  or  thin  beds  the  method  of  Fig.  4 
(a)  is  the  best.  Care  must  be  taken,  however,  in  the  case  of  beds  dipping 
over  45  deg.  to  stop  the  holes  of  the  horizontal  row  1  at  the  last  bedding 
plane  which  intersects  the  face  of  the  heading  above  the  floor. 

Case  VI. — Rocks  in  Inclined  Beds  Dipping  Away  from  the  Floor  of  the 
Heading. — For  either  medium  thick  or  thin  beds  the  method  of  Fig.  4  (c) 
should  be  used.  The  bar  is  set  up  at  A  for  row  2  and  at  B  for  rows  1  and  3. 
The  order  of  firing  the  horizontal  rows  of  holes  is  1,  2  and  finally  3. 
The  end  of  the  holes  in  row  1  should  be  stopped  beneath  the  last  bed- 
ding plane  intersecting  the  face  of  the  tunnel  under  the  roof  in  order 
to  utilize  this  plane  as  a  free  face  in  breaking. 

SURFACE  EXCAVATION  AND  UNDERGROUND  STOPING 

In  some  kinds  of  deposits,  especially  the  huge  copper-bearing  por- 
phyry bnses  and  the  Lake  Superior  iron  mines,  much  time  is  often 
saved  by  drilling  all  the  holes  possible  in  the  periphery  of  a  heading  in  the 
ore  from  the  same  set-ups  that  are  used  in  drilling  the  face.  These 
p3ripheral  holes  can  then  be  left  untouched  until  the  stoping  of  that 
ssction  begins,  when  they  can  be  easily  loaded  and  fired. 

Holes  for  stoping  may  be  placed  according  to  the  direction  in  three 
groups,  (1)  down  holes  (2)  flat-holes,  and  (3)  uppers.  A  dip  of  about 
45  deg.  downward  and  upward  can  be  assumed  to  make  the  limit 
between  groups  (1)  and  (2)  and  of  (2)  and  (3)  respectively,  although  the 
division  between  (2)  and  (3)  is  really  marked  by  the  angle  of  repose  of  the 
cuttings,  that  is,  when  the  hole  becomes  self-cleaning,  which  may  often 
mean  a  steeper  dip  than  45  deg.  The  speed  of  cutting  with  recipro- 
cating drills  depends  on  the  removal  of  cuttings  after  each  stroke  to 
expose  a  fresh  face.  Therefore  with  these  drills  down  holes  drill  easiest, 
then  uppers,  and  lastly  flats.  Using  the  hammer  drills  with  hollow  bits 
cleaned  by  water  or  air-jets,  there  is  less  difference  in  drilling  speed 
for  different  directions  of  pointing. 

Down  Holes;  Underground. — Down  holes  are  used  underground  in  the 
underhand  benches  of  tunnels  or  metal  mines.  To  start  this  system,  a 
heading  ah,  Fig.  8,  is  run  at  the  top  of  the  tunnel  or  stope  and  the  down 
holes  put  in  its  floor  for  the  first  bench.  The  depth  of  this  bench  is 
limited  by  the  length  of  the  bit  which  can  be  inserted  in  the  hole  and  that 
depends  on  the  height  of  the  heading  which  is  usually  around  7  ft.  so 
that  the  ordinary  railroad  tunnel,  20  to  25  ft.  high,  requires  two  benches 
and  two  settings  of  the  tripod  at  a  and  6,  Fig.  8  to  reach  the  bottom. 
These  bench  holes  point  downward  anyhow  but  often  an  advantage 
may  be  taken  of  the  structure.  Thus  with  horizontal  beds,  the  holes 
of  the  first  bench  can  be  terminated  at  a  bedding  plane  which  the  gases 


PRINCIPLES    OF   BLASTING    GROUND 


25 


from  the  explosion  will  enter  and  thus  exert  a  lifting  action  on  the  mass 
to  be  broken  off. 

Where  there  is  a  choice  of  plans,  a  heading  can  often  be  given  in  a 
direction  that  will  take  the  maximum  advantage  of  the  bedding  and 
joint  planes  for  breaking,  both  in  driving  and  stoping.  On  this  principle, 
the  rooms  of  coal  mines  are  usually  laid  put  perpendicular  to  the  line 
of  the  main  joint  planes  of  the  coal  seam  or  to  the  "face  cleat. " 

Down  Holes;  at  Surface. — Above  ground  the  only  limit  to  the  depth 
of  the  hole  is  the  capacity  of  the  drill.  In  considering  breaking  from 
deep  holes  we  have  a  choice  of  two  methods  (a)  multi-charging,  (6) 
chambering. 

In  the  drill  hole  ab  of  Fig.  9,  it  is  evident  that  a  charge  of  explosive 
at  any  point  b  will  only  break  out  a  cone  like  cbd  where  eb  is  the  line  of 


h     a 


v  ...     .;-.'.. 


f 


FIG.    8. — Holes  for  underhand  stope. 


FIG.  9. — Holes  for  high  bench. 


least  resistance.  In  order  to  break  the  whole  length  ab  by  multi-charg- 
ing, other  charges  of  explosives  as  /  and  g  would  be  placed  along  the  hole, 
with  tamping  between,  and  all  be  set  off  by  simultaneous  firing.  In  this 
way  the  whole  mass  abd  would  be  detached. 

By  chambering,  the  breaking  from  a  long  hole  would  be  achieved 
differently.  Instead  of  the  hole  being  placed  near  the  face  hd  of  the 
bench  as  is  the  hole  ab  (because  of  its  small  section  for  developing 
explosive  pressure),  the  hole  mn  would  be  placed  back  from  the  face  so 
that  nCj  the  line  of  least  resistance  in  homogeneous  rock,  would  be  only 
a  little  shorter  than  the  length  of  the  hole  above  the  chamber  at  n.  The 
chambering  is  effected  by  shattering  the  bottom  of  the  hole  with  high- 
power  dynamite  so  that  the  final  shape  of  the  chamber  approaches  a 
sphere.  In  France  this  chambering,  in  limestone,  is  performed  with 
hydro  shlor.'c  acip,  each  dose  of  neutralized  acid  being  washed  out  and  a 
new  one  poured  in  until  the  chamber  is  of  the  required  size.  When  the 
chamber  is  filled  with  gunpowder  or  low-power  dynamite  and  exploded, 


26 


MINING    WITHOUT   TIMBER 


it  will  exert  nearly  as  much  force  upward  as  horizontally  and  will  break 
out  a  mass  along  the  surface  of  fracture  qnp. 

The  choice  between  multi-charging  and  chambering  depends  on  the 
varying  conditions  of  formation,  drilling  and  exploding.  In  a  fissured 
formation,  chambering  has  often  an  advantage  because  the  explosive 
may  be  localized  in  a  solid  portion  of  the  rock,  although  it  often  needs  the 
use  of  two  kinds  of  explosives,  one  for  chambering  and  the  other  for 
breaking.  Where  it  is  desired  to  break  off  only  a  thin  slice  like  hab,  Fig. 
4,  from  the  cliff,  it  is  evident  that  multi-charging  should  be  resorted  to. 
When  an  even  topography  will  allow  the  handling  of  the  portable  steam 
or  electric  churn  drill  for  a3-in.  to  12-in.  hole  (instead  of  the  reciprocating 
drill  for  a  1  1/2-in.  hole),  the  multi-charging  method  will  permit  the 
drilling  and  breaking  of  a  much  longer  hole  than  would  be  feasible  by 
chambering. 

Flat  Holes;  Underground. — Of  the  three  groups,  flat  holes  are  the 
most  difficult  to  drill,  especially  those  which  are  pointed  above  the 
horizontal  for  the  reason  that  they  neither  hold  water  or  discharge  their 
cuttings  by  gravity.  This  group  is  much  used  in  the  overhead  stoping 
system  with  piston  drills  as  the  drill  tripod  can  be  set  on  the  broken 
rock.  Thus  flat  water  holes  which  are  easier  to  load  than  uppers  and 
free  from  their  dust  can  be  drilled  at  a  fair  speed.  In  overhand  stoping 
with  a  weak  back,  as  in  the  vertical  veins  of  Butte,  Mont.,  flat  holes 
have  also  an  advantage  over  uppers  as  the  timber  sets  can  be  carried  next 
to  the  back  and  the  drilling  can  proceed  under  the  lagging.  Thus  in 
Fig.  7  at  C  rows  1  and  2  are  water  holes  and  only  row  3  need  be  drilled  dry. 

In  the  zinc  district  of  Joplin,  Mo.,  flat  holes  are  used  instead  of  the 
usual  down  holes  to  break  the  benches  below  the  heading  of  the  under- 
hand stoping  system  as  described  under  Example  10  of  Chapter  VIII. 


I 

XV'^Vvv' 

/  /  ,/       x  \  N 

+  '        '   ff       /      /                                            \         S           US 

t  s  */   /a.                   c  \    ^v  n\  s 

Z 

IE 

—  r~ 

"1 

x                                                       y 

_L 

lio.  10. — Moles  tor  seam. 

In  driving  coal  headings  or  rooms  by  "blasting  off  the  solid/'  flat 
holes  bored  by  augers  are  generally  used  and  are  placed  similarly  to 
those  shown  for  headings  in  flatly  bedded  rock  in  Fig.  5.  In  the  location 
of  the  horizontal  rows  of  holes,  the  character  of  the  bedding  planes  be- 
tween the  coal  seam  and  its  roof  and  floor  must  be  considered.  If  the 
roof  is  "tight,"  the  shot  must  exert  a  strong  shearing  force  to  separate  it. 
This  is  achieved  by  slanting  the  row  of  holes  sharply  upward  and  ter- 


PRINCIPLES    OF    BLASTING    GROUND 


27 


minating  them  at  the  tight  plane.  A  similar  remedy  is  applied  to  a  tight 
floor.  In  many  seams  the  coal  is  cut  up  into  cubes  by  t\vo  sets  of  joint- 
planes  perpendicular  to  the  bedding  planes,  called  the  "face"  and  "end" 
cleats,  which  condition  makes  breaking  easy. 

The  shearing  of  a  coal  face,  before  shooting,  takes  the  place  of  the  cut 
holes  in  blasting  off  the  solid  and  the  smaller  charges  allowable  for  the 
former  method  not  only  save  explosive  but  prevent  the  shattering  of  the 
roof.  With  coal  sheared  vertically  along  one  rib  of  a  heading,  the  holes 
for  breaking  would  be  placed  like  vertical  rows  2  and  3,  Fig.  5.  Where 
the  shear  is  made  horizontally  as  in  the  undercut  xy,  Fig.  10,  it  is  custo- 
mary in  a  thick  seam  of  coal  to  place  the  first  or  "buster"  shot  at  6  in 
order  to  break  out  the  triangular  prism  of  coal  abc.  Then  when  the 
shattered  strip  gfh  has  been  removed  by  the  pick,  we  have  dm  and  en 
instead  of  dt  and  cs  for  the  line  of  least  resistance  from  the  corner  holes 
d  and  e,  by  which  last  the  balance  of  the  undercut  coal  can  now  be  easily 
shot  down.  For  a  thin  vein  of  coal,  the  "buster"  shot  would  be  located 
at  K  on  a  level  with  the  corner  holes  and  it  would 
break  out  the  triangular  prism  tKs  as  thick  as  the 
seam. 

The  undercut  shown  in  Fig.  10  is  that  made  by 
a  hand  or  power  pick.  Being  a  height  of  12  in.  or 
so  in  front  with  a  downward  slope  to  4  in.  in  the 
back,  its  shape  allows  the  "buster"  shot  to  throw 
much  of  the  coal  out  of  the  undercut,  so  that  the 
strip  gfh  can  be  easily  extracted  by  the  pick  to 
prepare  for  the  corner  shots.  When  the  undercut, 
however,  is  made  by  a  chain  machine,  it  is  of  uni- 
form height  of  only  about  4  in.,  and  the  "buster" 
shot  may  not  throw  the  coal  outward.  It  is  then 
often  advisable  to  place  an  extra  "snubbing"  shot 
at  /  to  flatten  down  the  detached  prism  abc  so  that 
the  shots  d  and  c  can  be  made  effective  without  first 
cleaning  out  the  broken  coal  underneath. 

Flat  Hole;  Surface. — In  loosening  huge  banks  of  placer  gravel  in 
California  before  hydraulicking,  small  adits  have  been  used  with  cross- 
cuts at  their  ends  to  hold  the  explosive.  From  a  breaking  stand-point, 
these  adits  correspond  to  flat  drill-holes  with  chambered  ends.  The 
same  method  has  also  been  employed  for  breaking  great  masses  of  rock 
in  quarries  or  excavations.  Often  a  shaft  has  been  sunk  as  an  entrance  to 
the  explosive  chamber  instead  of  an  adit.  Sometimes  two  cross-cuts 
from  the  adit  may  be  made  for  explosive  chambers,  as  shown  in  Fig.  11. 
There  only  the  crosscuts  cd  and  ab  would  be  packed  with  gunpowder  or 
low-power  dynamite,  while  the  adit  itself  would  be  blocked  with  timber 
or  masonry  bulkheads  wherever  it  met  the  crosscuts.  Elsewhere  it  would 


-50- 

Section 

FIG.  11. — Blasting  by 

tunnels. 


28  MINING    WITHOUT    TIMBER 

be  packed  with  sand.  For  firing,  electric  fuzes  or  caps  would  be  placed 
in  the  explosive  at  intervals  of  about  10  ft.  Finally  they  would  all  be 
connected  by  wiring  in  order  that  they  might  be  fired  simultaneously  by 
electricity  clk  being  the  line  of  least  resistance.  The  chamber  cd  would 
break  out  the  cone  gc1/1,  and  the  chamber  ab  would  break  out  the  prism 
halc^g,  the  plan  of  the  line  of  fracture  being  mabn. 

The  same  breaking  equation,  pa  =  TS,  applies  as  in  the  case  of  drill 
holes,  the  factor  a  being  the  area  of  the  cross  section  of  the  explosive 
taken  along  the  axis  of  the  crosscut. 

Uppers. — Uppers  are  seldom  used  on  the  surface  but  are  common  in 
underground  work  not  only  in  tunnel  headings  and  raises,  but  also  in 
overhand  stoping.  In  excavating  overhand  stopes  with  square-set 
timbering,  it  is  sometimes  more  efficient  to  drill  the  back  with  uppers 
as  at  B,  Fig.  7,  instead  of  the  flats  at  C  used  in  Butte  practice.  In  the 
great  stopes  of  the  Portland  mine,  at  Cripple  Creek,  Colo.,  where  the  pay 
shoot  was  in  places  120  ft.  wide  and  400  ft.  long,  the  ore  hard  and  the 
back  strong  enough  to  stay  up  across  the  vein  for  several  sets  ahead  of 
the  timbermen,  it  was  found  that  the  fastest  breaking  was  accomplished 
by  drill  ng  uppers  from  piston  drills  set  on  tripods,  one  drill  being  used 
in  every  set  across  the  stope. 


CHAPTER  III 
COMPRESSED  AIR  FOR  MINING 

In  drilling  with  piston  rock  drills  a  high  pressure  gives  a  stronger 
withdrawing  force  on  the  bit  which  tends  to  prevent  sticking  in  fissured 
ground  and  thus  greatly  increases  the  speed  of  boring.  In  hard,  tough 
ground,  like  specular  hematite  or  certain  intrusives,  a  high  air  pressure  is 
necessary,  if  it  is  desired  to  strike  a  blow,  severe  enough  to  cut  the  rock, 
with  a  light  portable  machine.  In  a  certain  mine,  using  40  drills  in  hard 
and  fissured  ground  the  rock  broken  per  machine  was  increased  about  20 
per  cent,  by  the  simple  expedient  of  advancing  the  air  pressure  from  75 
to  100  pounds.  A  low  pressure  system  requires  larger  pipes  to  deliver 
the  same  power  and  heavier  pumps  and  hoists  in  the  mine  to  accomplish 
a  given  amount  of  work  than  an  equivalent  equipment  working  under 
high  pressure. 

The  economical  limit  of  pressure  depends  in  a  given  case  on  commer- 
cial considerations,  costs  of  fuel,  labor  and  supplies,  which  in  turn  are 
governed  in  considerable  degree  by  the  mechanical  efficiency  of  the  plant. 
The  high  pressure  limit,  except  for  haulage  purposes  is  about  120  pounds. 


FIG.  12. — Relations  of  volume  and  pressure  in  air  compression. 

It  is  wasteful  to  heat  the  air  during  compression  to  a  higher  tempera- 
ture than  that  of  the  mine,  as  radiation  in  the  pipe-line  will  cool  any 
warmer  air  before  it  reaches  the  motor.  A  proof  of  this  statement 
follows:  Let  V  and  P  be  respectively  the  volume  and  pressure  of  free  air 
at  the  beginning  of  compression,  and  in  the  theoretical  indicator  card, 
Fig.  12,  in  which  0  is  the  origin  of  coordinates,  let  the  abscissa  of  point  a 
be  V  and  the  ordinate  be  P.  Let  V  and  P  be  the  volume  and  pressure 
of  air  at  any  point  of  the  stroke,  during  its  compression  by  a  reciprocating 
piston.  Then  if  the  temperature  due  to  the  heat  of  internal  friction  is 

29 


30  MINING    WITHOUT   TIMBER 

retained  in  the  air,  we  have  adiabatic  compression  and  get  the  curve 

a  b  /,  the  equation  of  which  is 

/  -    \ 


P  = 

the  value  of  y  being  1.406  for  dry  air  and  somewhat  less  for  the  ordinary' 
atmosphere,  and  p  being  the  resultant  pressure  and  v  the  resultant 
volume. 

If  the  temperature  is  kept  constant  during  compression,  by  removing 
the  internal  heat  as  fast  as  generated,  we  have  isothermal  compression 

and  get  the  cruve  a  cd,  the  equation  for  which  is  p!F»PY(—}.     Finally, 

\v/ 

the  work  lost  by  cooling  the  air,  from  the  final  adiabatic  temperature  to 
that  of  the  free  air,  is  measured  by  the  area  a  c  dfb,  the  total  work  of 
compression  for  one  stroke  of  the  piston  being  area  afmn. 

THEORY  OF  THE  INTERCOOLER 

Although  isothermal  compression  is  the  ideal,  practical  difficulties 
prevent  its  attainment.  The  air  can  be  cooled  in  the  compression  cylin- 
der by  a  water  spray,  but  this  method  requires  too  slow  a  machine  to 
compete  with  dry  compression  and  external  cooling.  It  can  be  easily 
shown,  mathematically  or  by  an  indicator  card,  that  water-jacketing 
the  compression  cylinder  has  practically  no  effect  in  cooling  the  air, 
although  it  is  useful  in  keeping  the  bearing  surfaces  cool  enough  for 
lubrication. 

In  Fig.  12,  the  adiabatic  and  isothermal  curves  get  farther  apart  as 
the  pressure  increases,  so  that  the  work  lost  by  adiabatic  compression 
increases  at  a  faster  ratio  than  the  pressure.  To  avoid  this  increase  for 
high  pressures,  a  compression  in  two  stages,  with  a  surface  intercooler 
between  the  high-  and  low-pressure  cylinders,  is  frequently  used.  Unfor- 
tunately, few  of  the  standard  machines  have  a  large  enough  intercooler 
to  insure  that  the  compressed  air,  entering  the  high-pressure  cylinder,  is 
as  cool  as  the  free  air  entering  the  low-pressure  cylinder  when  the  machine 
is  running  full  speed.  It  will  aid  the  intercooler,  if  the  free  air  is  sucked 
into  the  low-pressure  cylinder  from  the  coolest  available  place. 

In  the  diagram,  Fig.  12,  K  is  the  pressure  at  which  the  air  leaves  the 
low-pressure  cylinder  to  pass  through  the  intercooler  and  enter  the  high- 
pressure  cylinder.  The  following  cycle  then  takes  place  with  a  perfect 
intercooler.  In  the  low-pressure  cylinder  the  air  is  compressed  adia- 
batically  from  a  to  b,  reduced  in  the  intercooler  to  the  volume  at  point 
c  and  then  compressed  adiabatically  in  the  high-pressure  cylinder  from 
c  to  e,  the  total  work  of  compression  being  the  area  a  b  c  e  m  n.  Thus 
the  saving  of  work  by  the  use  of  the  intercooler  is  represented  by  the  area 
c  efb,  from  which  must  be  deducted  any  work  expended  in  circulating 


COMPRESSED    AIR    FOR    MINING  31 

the  cooling  water.  In  the  design  of  the  machine,  the  ratio  of  the  diam- 
eters of  the  low-pressure  cylinder  and  the  high-pressure  cylinder  are 
taken  so  that  the  area  a  b  k  n  is  equal  to  area  c  e  m  k  for  average 
conditions. 

There  need  be  little  difference  in  the  efficiency  of  the  steam  ends 
between  high-  and  low-pressure  compression.  With  a  cross-compound 
air  end,  the  steam  end  can  also  be  compound  and  for  a  single-stage  air 
end  the  machine  can  be  tandem-compound.  The  air-pressure  governor 
has  now  been  perfected  and  for  the  usual  variable  loads  of  mine  work, 
is  indispensable  for  any  pressure,  though  it  requires  a  duplex  machine 
to  avoid  a  stoppage  on  dead  center  with  no  load. 

PREHEATERS 

In  the  case  of  the  air  motor,  the  compression  process  is  reversed. 
The  air  on  entering  the  motor  in  the  mine  has  the  pressure  and  volume 
of  point  d  (Fig.  12)  and  in  a  simple,  unheated  motor  cylinder  will  expand 
adiabatically  along  the  line  d  g  h.  Should  the  air  be  preheated  to  the 
volume  of  point  /  it  will  then  expand  along  the  adiabatic  line  /  6  a  with  a 
gain  of  work,  over  the  unheated  case,  equal  to  area  a  h  df. 

With  two-stage  expansion,  the  air  may  be  preheated  before  entering 
the  low-pressure  cylinder  to  e,  then  expand  adiabatically  to  c,  next  pass 
through  an  interheater  so  as  to  reach  b  on  entering  the  high-pressure 
cylinder  and  finally  expand  adiabatically  to  a.  Heating  during  expan- 
sion, like  cooling  during  compression,  gains  in  its  relative  effect  on  the 
efficiency,  the  higher  the  pressure.  Aside  from  its  gain  in  work,  heating 
is  often  necessary  to  prevent  freezing  of  the  exhaust  when  the  air  is 
damp  and  cold  on  entering  the  motor. 

Owing  to  the  small  size  and  portability  of  rock  drills  preheaters  are 
for  this  service  out  of  place,  but  for  large  hoists  and  pumps,  with  high- 
pressure  air,  they  are  always  to  be  recommended.  In  the  operation  of 
the  -preheater  the  compressed  air  passes  through  a  vessel  containing 
heated  tubes  of  sufficient  radiating  surface  for  the  purpose.  These 
tubes  may  be  heated  by  a  coal,  coke  or  oil  fire,  but,  since  smoke 
contaminates  the  atmosphere  of  the  mine,  steam-heating  is  often  both 
convenient  and  economical.  In  an  air  heater  it  is  possible  to  utilize 
steam  more  efficiently  than  in  the  best  condensing  engine,  for  both 
the  latent  and  visible  heat  of  the  steam  are  absorbed  by  the  air 
and  turned  into  work  without  frictional  losses  greater  than  the  motor 
would  suffer  with  unheated  air.  With  steam  heating  the  only  im- 
portant loss  is  that  due  to  radiation  in  the  supply  pipe  from  the 
boilers,  and  by  proper  covering  this  can  be  made  small.  In  the  500-gal. 
Dickson  pumps,  installed  in  the  Anaconda  mines  at  Butte  in  1899,  the 
air  was  successfully  heated  by  steam  in  both  the  preheaters  and  the  inter- 
heaters  for  the  compound  cylinders. 


32  MINING    WITHOUT   TIMBER 

High-pressure  pipe-lines,  though  smaller  in  diameter,  require  more 
care  to  keep  them  tight  than  lines  for  low  pressure,  and  the  velocity  of 
exit  of  air  from  a  leak  varies  directly  as  the  square  of  the  pressure. 

The  loss  of  power  from  the  common  practice  of  blowing  out  powder 
smoke  with  the  air  hose  is  the  greater  the  higher  the  pressure,  for  the 
ventilating  efficiency  depends  only  on  the  quality  of  free  air  discharged. 
With  pipes  properly  proportioned  for  the  quantity  of  air  to  be  delivered 
the  frictional  line  losses  will  be  moderate  with  either  pressure,  if  care  be 
taken  to  avoid  unnecessary  bends  and  to  use  gate  valves  instead  of  globe 
valves. 

The  compressor  should  discharge  its  air  into  a  receiver  the  cooling 
action  of  which  will  not  only  at  cone  reduce  the  volume  to  that  which  it 
will  have  in  the  mine,  but  will  also  precipitate  any  extra  moisture  and 
keep  it  from  entering  the  pipe-lines.  A  good  device  for  the  surface 
receiver  is  a  condemned  boiler,  set  in  a  wooden  tank  in  which  is  water 
circulating  through  the  boiler  tubes,  while  the  compressed  air  fills  the 
shell.  Underground  the  receivers  need  only  be  plain  steel  shells  for 
storage,  but  they  must  be  numerous  and  large  enough  to  preserve  the 
pressure  constant  under  the  variable  power  requirements.  Preheaters 
in  use  serve  as  receivers. 

When  air  is  used  for  haulage  it  needs  a  special  piping  system  to  hold 
the  requisite  pressure  of  1000  Ibs.  upward.  This  piping  also  serves  as  a 
receiver  and  accumulator  of  air  between  locomotive  chargings  so  that 
the  compressors  can  be  run  under  a  constant  load.  It  is  evident 
that  the  piping  system  will  need  a  lesser  'proportionate  capacity 
as  receiver  the  greater  the  number  of  locomotives  supplied,  for 
each  charging  will  involve  a  less  relative  displacem3nt  of  air.  Under 
the  usual  traffic  and  air  pressure  a  pipe  line  of  6  to  12-in.  dia.  is  amply 
large,  both  for  distribution  and  storage  of  air,  without  placing  tank 
receivers  at  the  stations. 

The  air  ends  of  compressors  for  haulage  systems  should  be  at  least 
4-stage,  of  moderate  speed  and  with  ample  intercooling  surfaces;  for 
Fig.  12  shows  how  fast  the  power  loss  due  to  inefficient  cooling  increases 
with  the  pressure.  Until  recently  the  locomotives  were  single-stage  and 
had  consequently  a  low  efficiency  and  capacity;  but  the  new  compound, 
Porter  locomotive  obviates  these  troubles  and  gives  air-haulage  a  chance 
for  extension  beyond  its  present  special  field  of  gaseous  or  dusty  coal 
mines. 


CHAPTER  IV 
PRINCIPLES  FOR  CONTROLLING  EXCAVATIONS 

The  art  of  timbering  is  not  synonymous  with  that  of  the  control  of 
ground  as  some  suppose;  a  good  carpenter  can  frame  timber  better  than 
any  miner,  but  unless  he  places  it  underground  as  directed  by  the  latter, 
his  accurately  jointed  sets  are  liable  to  prove  worthless  for  the  purpose 
intended.  The  subject  of  ground  control  naturally  divides  itself  under 
two  topics:  I.  The  control  of  the  roof  of  an  excavation;  II.  The  control 
of  the  sides  and  floor  of  an  excavation  and  of  the  whole  overlying  forma- 
tion. Both  topics  will  be  considered  separately  before  their  inter-relation 
will  be  discussed.  In  practice,  we  have  not  only  to  consider  the 
freshly  broken  surfaces  of  an  excavation,  but  their  future  conditions  after 
exposure  to  the  weathering  action  of  the  mine  atmosphere. 

CONTROL  OF  THE  ROOF 

(a)  Roof  over  a  Horizontal  Room. — This  case  is  the  simplest  and 
occurs  in  mining  horizontal  seams  or  beds.  Let  abb'a',  Fig.  13,  represent 
the  cross-section  of  a  rectangular  room  excavated  in  a  team  of  the  thick- 


FIG.   13. — Homogeneous  or  horizontally-bedded  roof. 

ness  aa' '.  Then  the  support  of  the  roof  over  the  opening  ab  depends  upon 
the  immediately  overlying  formation.  The  structure  of  the  last  falls 
usually  under  one  of  the  five  following  cases:  (1)  homogeneous,  (2) 
horizontally  bedded,  (3)  weakly-consolidated,  (4)  non-conformable,  (5) 
broken. 

With  case  (1)  or  a  homogeneous  roof  stratum,  either  massive  or  in  a 
sufficiently  thick  bed  to  act  as  such,  the  lines  of  vertical  pressure  far 
3  33 


34  MINING    WITHOUT    TIMBER 

above  ab  tend  to  combine  themselves  into  resultants  which  follow  a 
surface  acb  and  throw  the  downward  pressure  onto  the  walls  at  a  and  6. 
The  resultant  surface  takes  the  form  of  an  arch,  over  a  tunnel,  or  of  an 
arch  with  domed  ends,  in  a  room  of  limited  length.  This  means  that  the 
sub-arch  block  acb  is  all  the  weight  that  has  to  be  supported  to  maintain 
the  roof  intact,  and  that  its  stability  depends  first,  on  its  strength  as  a 
beam  of  continuous  width  to  bear  its  own  weight  across  the  span  ab  and 
second,  on  its  being  held  in  place  by  the  tensile  strength  of  the  rock  area 
along  the  arch,  or  potential  surface  of  fracture,  acb.  In  case  (1)  the 
natural  arching  is  usually  sufficiently  convex  so  that  the  sub-arch-block 
has  sufficient  depth  cf  to  make  is  self-sustaining  as  a  beam  across  ab 
except  in  soft  rocks  like  certain  shales  which  may  not  only  need  the  sup- 
port of  a  cap  like  ab  but  also  must  be  lagged. 

In  old  mine  workings  where  the  sub-arch  block  acb  has  fallen  out  so 
that  the  shape  of  the  natural  surface  of  equilibrium  acb  can  be  discerned, 
it  appears  as  an  arch  whose  proportions  vary  with  the  width  of  the  room 
and  the  nature  of  the  roof.  Fayal  gives  as  working  rules  for  limited 
areas  like  rooms: 

If  w  =  width  of  room  (as  ab  in  Fig.  13) ; 
h  =  height  of  arch  (as  c/in  Fig.  13); 

If  w  is  less  than  6  ft.,  h  may  be  as  much  as  2w  (Fayol's  first  rule) ; 

If  w  is  more  than  6  ft.,  h  may  be  as  much  as  4w  (Fayol's  second  rule). 

In  railroad  or  mine  tunnels,  a  homogeneous  roof  can  be  made  self-sus- 
taining by  excavating  it,  at  the  start,  along  the  natural  arch  form.  In 
the  rooms  of  coal  seams,  however,  or  in  iron-ore  beds,  the  sub-arch  block 
must  be  sustained  intact  until  the  mineral  beneath  is  removed  and  the 
room  abandoned.  The  tensile  strength  of  the  arched  surface  is  seldom 
sufficient  to  accomplish  this  unaided,  except  in  narrow  rooms.  In  wider 
rooms,  a  cross-beam  ab,  or  one  or  more  props  like  ff  must  be  put  in  whose 
strength,  however,  need  only  equal  the  difference  between  the  weight  of 
the  sub-arch  block  acb  and  the  tensile  strength  of  the  arched  surface  acb, 
provided  that  ff  is  inserted  before  the  surface  acb  has  begun  to  fracture. 
Should  the  latter  accident  have  taken  place,  the  weight  of  the  whole  sub- 
arch  block  may  have  to  be  sustained  by  props  and  thus  a  heavy  unneces- 
sary expense  be  incurred. 

With  case  (2)  or  where  the  roof  is  in  beds  thinner  than  the  sub-arch 
block  so  that  bedding  planes  like  hk  and  mn  (Fig.  13)  intersect  the  surface 
acb,  a  different  condition  arises  from  case  (1).  It  is  evident  that  now 
the  sub-arch  block  instead  of  being  a  single  stone  beam  acb  is  divided 
into  three  stone  beams  ahkb,  h'mnk'  and  m'dd'n' ' ,  so  that  for  a  self-sustain- 
ing roof,  the  lowest  beam  ahkb  must  be  strong  enough  to  sustain  the 
weight  of  the  two  beams  above  it,  the  central  beam  h'mnk'  must  sustain 
the  top  beam  m'dd'n'  and  the  tensile  strength  of  the  sub-arch  surface  acb 
must  be,  as  before,  sufficiently  strong  to  hold  up  the  whole  sub-arch 


PRINCIPLES    FOR    CONTROLLING    EXCAVATIONS  35 

block.  It  is,  therefore,  likely  that  a  room  would  need  stronger  props  in 
case  (2)  than  in  case  (1)  because  the  lowest  sustaining  beam  of  case  (2) 
has  a  depth  at  the  middle  of  ha,  which  is  only  a  fraction  of  the  correspond- 
ing depth  cf  for  case  (1),  and  the  cross  breaking  strength  of  a  beam 
increases  directly  as  the  square  of  the  depth.  Also  we  now  do  not  have 
a  uniform  tensile  strength  for  one  surface  of  fracture  acb,  but  a  different 
strength  for  each  of  the  three  beds  which  acb  intersects.  Hence  for  case 
(2)  we  have  to  acsertain  both  the  cross  breaking  and  the  tensile  strengths 
of  all  beds  in  the  sub-arch  block  before  we  can  ascertain  how  much  prop- 
ping is  required  to  sustain  the  roof  across  a  room  of  a  given  width.  A 
roof  of  an  elastic  nature  like  slate  may  at  first  simply  bow  downward 
from  an  excess  of  pressure  instead  of  fracturing  as  a  beam.  This  may 
cause  it  to  fail  by  shear  at  the  abutments.  For  the  maximum  strength 
of  a  roof  it  is  important  to  exclude  water  from  the  bedding  planes  in  order 
to  prevent  the  slipping  and  weakness  caused  by  its  presence. 

Case  (3)  often  occurs  in  coal  mines  where  the  roof  stratum  is  "  clod  " 
or  a  kind  of  soft  shale  containing  concretions  of  considerable  size.  A 
common  device  is  to  leave  the  upper  layer  of  the  coal  seam  under  it  which 
then  acts  as  the  lowest  beam  ahkb  of  case  (2)  to  partially  sustain  the 
clod-stratum.  Where  all  the  seam  must  be  removed  beneath  the  clod, 
the  roof  can  only  be  kept  intact  by  excavating  the  mineral  with  little  or 
no  blasting  and  keeping  the  supports  close  to  the  working  face.  Props, 
cross-pieces  and  lagging  may  all  have  to  be  used.  If  the  clod  stratum  is 
thicker  than  the  room's  natural  arch,  masses  may  fall  out  from  above 
the  surface  acb,  after  the  sub-arch  block  has  been  taken  down,  so  that 
roofs  must  be  arched  higher  than  the  clod  in  order  to  stand  permanently 
unsupported. 

Where  the  roof  is  a  weakly  consolidated  stratum  of  more  uniformly 
sized  stones,  like  a  conglomerate,  the  problem  of  support  is  similar. 
Practically  the  whole  weight  of  the  sub-arch  block  must  be  held  in  place 
by  artificial  supports,  and  in  addition  the  beam  acb  itself  must  be  rein- 
forced by  cross-beams  of  props  or  by  both.  Where  the  roof  stratum  is 
so  weakly  consolidated  as  to  be  incoherent  it  requires  close  lagging,  and 
where  quite  loose  an  advance  can  only  be  made  by  driving  fore-poles 
ahead  of  the  timber  sets  right  up  to  the  working  face.  Loose  sand,  if 
dry  or  only  moist,  can  be  sustained,  like  loose  gravel,  by  close  fore-poling, 
but  if  it  is  wet  enough  to  flow  freely  like  quicksand  the  case  is  hopeless 
except  by  the  use  of  some  such  system  as  that  of  the  pneumatic  shield 
recently  employed  in  the  Hudson  river  tunnels  at  New  York. 

It  is  evident,  however,  that  while  the  quicksand  roof  of  a  railroad 
tunnel  might  be  penetrated  and  sustained  by  the  expensive  pneumatic 
shield  and  its  follower,  a  cast-iron  tube  lining,  such  a  device  would  be 
commercially  unpractical  for  ordinary  ore  deposits.  For  the  latter  the 
only  hope  for  overcoming  quicksand  is  sufficient  drainage  so  that  the 


36 


MINING    WITHOUT    TIMBER. 


sand  loses  its  fluidity  and  takes  the  compact  condition  of  its  merely  moist 
state.  If  drainage  of  the  quicksand  covering  is  not  feasible  and  the  ore 
body  cannot  be  mined  by  some  subaqueous  method,  it  is  worthless,  as 
was  recently  proved  for  a  huge  hematite  deposit  under  a  swamp  on  the 
Mesabi  range,  Minn.,  which  was  abandoned  after  wasting  a  large  suni  in 
attempting  to  open  a  mine  in  it. 

Even  if  the  bed  or  pocket  of  quicksand  does  not  rest  directly  on  the 
ore  body,  but  is  separated  from  it  by  a  rock  stratum,  great  care  has  to  be 
taken  against  it.  The  only  safe  plan  is  to  open  the  mine  excavations 
of  small  size  and  with  sufficient  support  to  keep  the  rock  roof  intact,  for 
otherwise  vertical  cracks  may  develop  reaching  to  the  quicksand. 
When  the  quicksand  once  begins  to  flow  into  the  mine,  the  results  may 


.;.:j::.':.63 
3  re  '••-.'• 

,•'••••••-'•'•, 
r 

f                 f' 

."••-•     " 

s 

3-.  ..;•.:.••..-.:.••... 

'"YvX^vo^^'sr'i 

Fia.  14. — Non-conformable  roof. 


be  far-reaching  for  its  escape  from  its  matrix  may  mean  the  collapse  of 
the  latter  and  consequent  disastrous  movements  of  the  whole  overlying 
cover. 

In  case  (4)  the  non-conformable  roof  strata  may  dip  in  any  direction 
with  reference  to  the  underlying  mineral  seam.  If  the  roof  strata  strike 
along  the  long  axis  of  the  room  as  in  the  cross  section  of  Fig.  14,  then  it  is 
evident  that  conditions  will  not  produce  an  arch  of  fracture  as  in  Fig.  13. 
The  upper  stratum  g'gpv  is  entirely  above  the  room  opening  and  bridges 
it  slantingly  from  one  side  to  the  other,  while  the  two  lower  strata,  vpv'q' 
and  q'v'q,  have  their  lower  ends  unsupported  and  projecting  like  canti- 
lever beams.  Then  the  natural  surfaces  of  fracture  will  be  normal  to  the 
bedding  planes  and  will  be  qq'  for  the  lowest  and  q'v  for  the  middle  stratum. 
The  tensile  strength  of  surface  qq'  must  be  enough  to  hold  the  weight  of 
projection  q'v'q  and  any  unbalanced  pressure  from  above,  while  surface 
vq'  must  hold  its  end  vpv'q'  and  any  weight  above.  A  line  of  props  at  t 
strong  enough  to  sustain  the  excess  of  strain  over  the  resisting  strength 
of  surface  qq' ,  will  hold  up  the  roof  without  a  second  line  at  t  provided 
that  surface  v'p'  is  strong  enough  to  sustain  the  weight  on  it  from  the 
projection  p'pv'. 


PRINCIPLES    FOR   CONTROLLING    EXCAVATIONS 


37 


As  the  dip  of  the  roof  beds  increases,  the  strain  on  the  surface  of 
fracture  qq'v  becomes  more  tensile  than  cross-breaking  until  with  vertical 
beds  the  strain  is  all  tensile.  In  the  last  case,  the  weight  to  be  sustained 
by  each  stratum  is  its  block  below  a  natural  arch  of  fracture  across  the 
room,  which  is  differently  proportioned  for  vertical  beds  than  is  acb  of 
Fig.  13  for  horizontal  beds. 

Where  the  strike  of  the  inclined  beds  of  the  roof  is  across  instead  of 
along  the  room  beneath,  we  have  a  mixture  of  the  cases  illustrated  by 
Figs.  13  and  14.  Each  bed  can  first  be  considered  separately  as  forming 
a  single  sub-arch  beam  whose  side  elevation  is  acb  in  Fig.  13.  Each  bed 
must  then  be  calculated  separately  both  for  the  self-sustaining  power  of 
its  sub-arch  beam,  across  the  span  of  the  room,  and  for  that  of  its  tensile 
surface  acb.  A  bed  may  then  be  artificially  supported  if  necessary,  by 
prop  ff  or  cap  ab..t  If  Fig.  14  be  assumed,  for  this  case  only,  to  be  the 
longitudinal  section  of  the  room  whose  cross-section  is  Fig.  13,  we  see  that 
a  cross  bed  like  vpv'q'm&y  have  the  same  breaking-off  action  on  a  lower 
bed  q'v'q  as  has  just  been  discussed  in  the  last  paragraph,  and  supports 
must  be  modified  accordingly. 

The  broken  roof  of  case  (5)  may  arise  from  planes  of  faulting,  fractur- 
ng,  jointing,  etc.  If  the  breaking  planes  are  parallel  or  r$  one  general 


FIG.  15. — Roof-over  inclined  room. 

direction,  we  can  handle  the  roof  as  suggested  for  case  (4).  If  the  planes 
are  in  several  directions  so  as  to  cut  the  roof  into  monoliths,  the  support 
of  each  block  will  have  to  be  studied  separately.  Where  a  roof  monolith 
is  of  indefinite  height,  we  may  illustrate  it  by  Fig.  13  with  ab  its  length 
and  acb  the  section  of  its  natural  surface  of  fracture,  which  will  be  of 
dome  shape,  so  that  only  the  support  of  the  sub-arch  portion  acb  has  then 
to  be  considered.  When,  however,  the  roof  monoliths  are  broken  also 
by  a  plane  in  a  horizontal  direction,  like  mn  in  Fig.  13,  so  as  to  become 


38  MINING   WITHOUT   TIMBER 

free  blocks  like  amnb,  they  can  only  be  kept  in  the  roof  by  sustaining 
their  entire  weight  artificially,  and  fore-poling  will  have  to  be  used  for 
excavating  beneath  a  roof  surface  containing  them. 

(6)  Roof  Over  an  Inclined  Room. — This  case  occurs  in  mining  seams  on 
a  dip  which  may  vary  up  to  90  deg.  from  the  horizontal.  Let  abb' a',  Fig.  15, 
represent  the  cross-section  of  a  room  in  a  seam  of  the  thickness  66',  which 
has  the  usual  horizontal  floor  aa'  for  tramming.  It  is  evident  that  the 
principles  of  roof  support  similar  to  the  previous  case  of  horizontal  rooms 
apply  here,  but  the  action  of  the  superincumbent  weight  in  the  roof  is 
affected  by  the  angle  of  dip.  Thus  in  the  diagram  of  Fig.  15,  if  W  =  super- 
incumbent weight;  6  =  angle  of  dip;  N  =  normal  pressure  on  roof; 
T  =  tangential  pressure  on  roof;  then 

N=W  cosd (1) 

and  T=  W  sin  6 (2) 

For  homogeneous  strata  the  weight  of  the  overlying  formation  would 
be  thrown  onto  the  pillars  at  a  and  b  and  the  potential  surface  of  fracture 
would  be  the  arch  acb.  Thus  the  span  ab  has  only  to  sustain  the  normal 
pressure  of  the  sub-arch  block  acb  acting  both  in  tension  on  the  surface 
acb  and  in  cross-breaking  strain  on  the  beam  acb  as  described  for  case  (a) 
in  Fig.  13.  The  back  of  the  ore  should  also  fall  on  the  arch  line  bdbf 
instead  of  a  straight  line  from  b  to  b'.  A  prop  to  hold  up  the  roof  will 
be  subjected  to  the  least  pressure  and  be  of  shortest  length  if  it  is  placed 
in  a  line  gfe  drawn  normal  to  the  hanging  wall  from  the  center  of  gravity 
of  the  sub-arch  block  at  g.  Because  of  possible  shrinkage  of  prop  or 
movement  of  ground,  however,  which  would  cause  a  normal  prop  to  fall 
out,  the  usual  practice  is  to  incline  it  about  10  deg.  downward  from  the 
normal  line  as  ff.  The  sub-arch  block  acb  can  also  be  sustained  by  a 
cap  ab  from  the  back  to  the  floor,  or  by  both  prop  and  cap.  With  the 
roof-strata  bedded  parallel  to  the  seam,  the  surface  of  fracture  assumes 
the  stepped-arch  form  ahh'mm'k'kb.  In  comparing  the  strains  and  the 
support  of  the  bedded  roof,  as  well  as  of  the  weakly-consolidated,  o.r  the 
non-conformable  and  the  broken  roofs  with  those  of  the  homogeneous 
roof,  the  same  differences  arise  as  already  explained  for  case  (a) . 

CONTROL  OF  THE  OVERLYING  FORMATIONS 

It  is  evident  that  when  part  of  a  bed  is  removed,  the  balance  left  as 
pillars  must  sustain  the  whole  overlying  formation.  There  are  three 
factors  that  enter  into  pillar  calculations,  the  roof,  the  pillars  or  sides 
and  the  floor.  The  stability  of  the  room  does  not  depend  alone  on  the 
strength  of  the  pillars  as  columns  for  an  excess  of  pressure  may  force  a 
sound  pillar  into  a  roof  or  floor  of  insufficient  compressive  strength  and 
cause  a  settling.  This  happens  with  materials  like  clay  which  are  hard 


PRINCIPLES    FOR    CONTROLLING    EXCAVATIONS  39 

when  dry  and  become  soft  when  moist,  so  unless  they  can  be  kept  dry 
during  mining,  the  pillar  calculations  must  guard  against  their  moist 
state.  An  excess  of  pressure  on  a  plastic  floor  will  cause  it  to  spread 
laterally  and  rise  from  under  the  lower  periphery  of  the  pillar,  thus 
exerting  a  horizontal  rending  force  on  the  latter  which  tends  to  disrupt 
its  edges. 

Any  downward  bowing  of  an  elastic  roof  over  the  rooms  must  be 
compensated  for  by  an  upward  bowing  over  the  interior  of  the  pillars. 
This  causes  an  oblique  pressure  at  the  upper  edges  of  the  latter  which 
tends  to  shear  them  off  as  the  roof  bends  more  and  more.  The  obliquity 
of  this  roof  pressure  on  the  pillar  edges  is  also  often  increased  by  a  rolling 
floor. 

Thus  a  mine  floor  and  roof  act  not  only  vertically  on  the  areas  of 
contact  with  the  pillars,  but  also  laterally,  while  the  bowing  of  the  roof 
produces  strains  parallel  to  the  strata  that  tend  to  separate  them  along 
their  bedding  planes,  and  thus  weaken  the  cross  breaking  strength  of  the 
roof.  For  these  reasons,  a  mine  pillar  will  stand  best  and  can  be  made  of 
the  minimum  volume  when  its  base  and  capital  meet  floor  and  roof  in 
broadly  spreading  tangential  curves,  which  are  concave  in  profile. 

Sometimes  the  roof  and  floor  beds,  in  direct  contact  with  the  seam, 
are  themselves  quite  hard,  but  so  thin  that  they  bend  and  transmit  the 
pressure  on  the  pillars  to  an  adjoining  soft  stratum  and  force  it  out 
through  any  fissures  that  may  be  in  the  roof  or  floor.  When  the  pillars 
themselves  are  too  weak  for  the  pressure,  what  is  called  a  "squeeze" 
(failure  of  pillars)  begins,  by  a  shelling  off  of  the  outer  surface,  and  later 
a  collapse  occurs,  which  may  be  a  gradual  sinking,  with  elastic  strata 
like  some  coals  and  shales  or  a  sudden  fracturing  in  masses  with  hard 
blocky  rock-like  limestone  or  quartz.  Some  substances,  like  coal,  pyrite 
and  easily  weathered  rocks,  loose  strength  on  exposure  to  mine  air  and 
this  fact  must  be  considered  if  durable  pillars  are  to  be  made  of  them. 
The  "minimum  fraction  of  a  bed  necessary  to  leave  for  pillars  may  be 
thus  calculated: 

Let  x    =  fraction  of  area  to  be  left  in  pillars; 
h  =  depth  of  cover  in  feet; 

w   =  specific  weight  of  cover  in  pounds  per  cu.  ft.; 
s   =  ultimate  compressive  strength  in  pounds  per  square  foot  of 

least  resistant  stratum  adjoining  pillar; 
m  =  factor  of  safety, 

Then,  weight  held  by  1  sq.  ft.  of  seam  =  hw,  and  compressive  strength 
of  corresponding  pillar  =  xms,  hence  hw  —  xms. 

>   hw 

or  x  =  — (3) 

ms 

If  excavation  is  inclined  at  an  angle  as  in  Fig.  15,  then  the  pressure  is 


40  MINING    WITHOUT    TIMBER 

hw  cos  0 

hwcosu,  so  that  hw  cos  6  —  xms  or  x  =  — (4) 

ms 

It  is  difficult  to  get  the  real  strength  of  the  floor,  pillar  and  roof  beds 
because  the  beds  themselves  are  seldom  free  from  planes  of  weakness 
which  would  not  be  appreciable  in  the  small  blocks  that  must  be  used  in 
the  testing  machines  for  compression,  tension  or  shear.  For  this  reason 
the  factor  of  safety  m  of  equations  (3)  and  (4)  is  taken  at  from  2  to  10, 
varying  with  the  nature  of  the  strata  and  of  the  mine  layout. 

It  is  only  by  close  watching  on  the  changing  conditions  that  move- 
ments of  the  formation  over  wide  excavations  can  be  prevented  even  in 
well  laid  out  mines.  An  incipient  "squeeze"  of  pillars  may  sometimes 
be  checked  by  building  up  stone-filled  wooden  cribs  along  their  edges, 
but  this  remedy  may  merely  shift  the  pressure  and  transfer  the  "  squeeze  " 
elsewhere.  Often  it  is  better  to  localize  rather  than  to  attempt  to  support 
a  squeeze  and  this  can  be  affected  by  allowing  the  roof  to  cave  over  the 
disturbed  section,  assisting  the  fall  where  necessary  by  blasting  the  roof 
and  pillars.  The  volume  of  roof  thus  made  to  fall  will  be  that  under  the 
dome  of  fracture  as  acb  of  Fig.  13,  the  span  ab  in -this  case  not  being  the 
width  of  a  single  room,  but  of  the  whole  disturbed  section.  If  the  seam 
is  thin  in  proportion  to  the  height  of  the  falling  dome,  the  broken  rock, 
as  it  occupies  more  space  than  when  solid,  will  fill  up  the  space  under  the 
surface  of  fracture  and  form  a  sufficient  support  to  prevent  any  further 
strain  on  the  overlying  formation. 

The  caving  of  the  roof  over  the  disturbed  area  is  also  a  remedy  for 
"creep"  (oozing  of  roof  or  floor  into  excavations),  but  if  the  ground 
surface  is  to  remain  intact  a  safer  plan  is  to  fill  the  excavation  solid  with 
rock.  Where  a  supply  of  fine  material  like  mill  tailing  or  sand  can  be 
obtained  cheaply,  the  filling  is  best  done  by  mixing  it  with  water  and 
running  it  into  the  workings  through  pipes  by  the  flushing  system  of  the 
Pennsylvania  anthracite  regions  as  described  in  Examples  58  and  59.  ' 

The  caving  of  the  roof,  locally,  by  blasting  can  be  easiest  affected  by 
reversing  the  methods  already  explained  for  roof  support.  If  pulling  or 
blasting  out  all  artificial  supports  does  not  bring  down  the  roof,  any  rock 
pillars  in  the  area  should  be  drilled  and  blasted  by  simultaneous  firing. 
The  next  procedure  is  to  drill  holes  into  the  roof  so  as  to  cut  a  groove 
around  the  springing  line  of  the  dome  acb  in  Fig.  13.  The  work  of  the  drill 
men  around  the  edges  of  the  excavation  will  be  safe  and  the  circumferen- 
tial groove  can  thus  easily  be  widened  and  carried  higher  until  the  central 
bell  of  the  sub-arch  dome  has  so  much  of  its  sustaining  surface  acb  cut 
away  that  it  drops  out. 

EFFECT  OF  CAVING  ON  OVERLYING  OBJECTS  , 

In  working  superimposed  beds  simultaneously,  it  is  necessary  to 
determine  the  proper  relative  position  of  pillars  in  the  various  beds. 


PRINCIPLES    FOR    CONTROLLING    EXCAVATIONS 


41 


Pillars  must  also  be  located,  in  caving  mines,  where  it  is  desired  to  pro- 
tect valuable  surface  structures.  In  modern  coal  mining,  both  the  long- 
wall  and  usually  the  room  and  pillar  method  involve  the  caving  of  the 
excavations. 

How  far  up  an  underground  subsidence  will  reach  depends  on  a  num- 
ber of  conditions,  such  as  area,  height  and  manner  of  making  of  excava- 
tion, nature  of  overlying  formation,  presence  of  faults  and  dikes,  etc.  By 
Fayol's  second  rule,  the  height- affected  by  subsidence  would  not  exceed 
four  times  the  width  of  the  excavation,  but  this  only  holds  good  for  a 
limited  area  whose  sub-arch  roof  block  can  scale  off  at  leisure.  When 
large  areas  are  excavated,  complex  stresses  arise  which  are  apt  to  cause 
sudden  irresistible  strains  on  the  roof  which  cause  it  to  develop  long  cracks 
and  fractures  analogous  to  faults.  If  the  overlying  strata  contain  many 
strong  rock  beds,  these  may  act  as  beams  which  rest  on  the  broken  caved 
formation  beneath  them  and  prevent  any  effect  above.  Thus  at  Sunder- 
land,  England,  where  half  of  the  strata  are  hard  rock,  coal  seams  have 


FIG.  16. — Effect  of  excavation  on  overlying  bed  and  on  surface. 

been  mined  and  caved  at  the  depths  of  1600  ft.  without  affecting  the 
surface.  In  the  Transvaal  gold  beds,  dipping  at  around  40  deg.,  caves  may 
occur  over  areas  of  several  acres  at  depths  over  1000  ft.  without  surface 
movement.  With  a  formation  of  soft  friable  strata,  like  shale  or  glacial 
drift,  however,  there  is  nothing  to  arrest  a  subsidence  beneath,  and  under 
such  roofs  the  effect  of  caving  coal  mines,  2000  ft.  deep,  has  depressed 
surface  structures. 

Fayal's  third  rule  applies  to  excavations  of  large  area  and  is  "  where 
the  area  is  infinite  and  the  beds  are  chiefly  sandstone  with  a  dip  less  than 


42  MINING    WITHOUT    TIMBEK 

40  deg.,  the  height  of  the  zone  of  subsidence  is  less  than  200  times  the 
height  of  the  excavation."  This  means  that  the  caving  of  an  excavation, 
6  ft.  high,  would  not  affect  the  surface  if  over  1200  ft.  be'ow  it.  The  third 
rule  is  based  on  the  height  of  excavation  rather  than  on  its  width,  like  the 
other  rules,  and  depends  on  the  principle  already  mentioned  that  the 
strong  strata  tend  to  rest  solidly,  ultimately,  on  the  caved  ground  below. 
Subsidence  does  not  break  strata  perpendicular  to  their  bedding 
planes.  For  defining  the  disturbed  area  over  excavations  under  unbroken 
stratified  formations  two  rules  are  used,  the  first  for  slightly  and  the 
second  for  steeply  dipping  roofs.  Thus  in  Fig.  16, 
if  D  =  dip  of  roof  strata  in  degrees 

A   =  dip  of  angle  of  fracture, 
for  roofs  under  30  deg.  dip  Richardson  gives, 

A  =90  deg.- 1/2  D (5) 

which  signifies  that  the  plane  of  fracture  e  f  (Fig.  16)  of  bed  ab  lies  half 
way  between  the  vertical  and  the  plane  eg  (normal  to  the  dip  line  of  the 
roof) . 

For  roofs  over  30  deg.  dip  Hausse  gives, 

tan  A  =  cotan  2D  +3  cosec  2D (6) 

Formula  (6)  gives  for  a30-deg.  roof  only  a  slightly  larger  angle  of  fracture 
than  formula  (5),  but  as  the  dip  gets  steeper  the  difference  between  the 
two  formulas  steadily  increases  while  a  maximum  A  is  reached  with 
formula  (6)  when  D  is  between  50  deg.  and  60  deg.  as  shown  in  the 
following  table: 


Dip  D 

0  deg. 
10  deg. 
20  deg. 
30  deg. 
40  deg. 
50  deg. 
60  deg. 
70  deg. 
80  deg. 
90  deg. 

These  formulae  can  only  be  considered  as  general  guides  to  the  prob- 
able location  of  the  plane  of  fracture  and  they  must  be  modified  in  practice 
by  a  consideration  of  the  surface  topography,  of  the  structure  of  the 
formation  and  of  natural  breaks  like  joints  and  faults.  Where  thick 
dikes  cut  across  the  roof  strata,  the  plane  of  fracture  is  more  apt  to  follow 
along  the  surface  of  the  dike  than  to  break  it. 


Angle  A,  d< 

3gS. 

Formula  (5) 

Formula  (6) 

90 

90.0 

85 

85.2 

80 

80.5 

75 

76.2 

70 

73.0 

65 

70.8 

60 

71.0 

55 

74.0 

50 

80.8 

45 

90.0 

PRINCIPLES    FOR    CONTROLLING    EXCAVATIONS  43 

Protection  of  Surface. — The  practical  use  of  these  formula  is  shown 
in  Fig.  16  where  it  is  desired  to  protect  the  building  at  /&'  when  mining 
the  veins  ab  and  cd.  Here  h'f  and  hf  are  planes  drawn  parallel  to  the 
plane  of  fracture  ef  and  their  intersection  with  the  beds  defines  the  inside 
boundaries  of  the  pillars  e'e  and  h'h.  The  margin  of  safety  to  be  left 
around  these  inside  limits  of  the  pillars  for  "draw"  varies  with  the 
importance  of  the  building  and  how  closely  the  strata  have  been  observed 
to  follow  the  fracturing  formulae. 

Protection  of  Overlying  Beds. — Where  the  veins  cd  and  ab  are  being 
mined  simultaneous' y,  it  is  evident  that  the  pillars  to  be  left  in  cd  to 
protect  pillar  e'e  must  not  be  the  ground  vertically  beneath,  as  hk,  but 
that  enclosed  between  the  same  planes  of  roof  fracture  as  h'h  with  a  due 
allowance  added  for  "  draw."  In  excavating,  also,  the  direction  of  the 
roof  fracture  ef  must  be  taken  instead  of  the  vertical  plane  as  the  guide 
to  relative  operations  in  the  upper  and  the  lower  beds.  Thus  for  safety 
the  bed  ab  would  be  stopped  "ahead"  (measuring  from  the  plane  ef)  of 
bed  cd;  except  in  the  case  where  cd  was  being  filled,  when  the  slight  sub- 
sidence of  the  floor  of  ab,  caused  by  the  settling  of  cd  (when  "  ahead  ")  on 
its  filling,  would  render  the  breaking  of  ab  easier. 

In  mining  the  superincumbent  parallel  anthracite  seams  of  the  Lehigh 
Valley  Coal  Co.,  by  the  room  and  pillar  system,  the  pillars  must  overlay 
each  other  when  the  parting  is  thin.  A  neglect  of  this  precaution,  with 
the  usual  parting,  is  liable  to  result  in  the  squeezing  of  the  overlying  pillars 
down  into  the  rooms  of  the  seam  below.  When  the  parting  is  over  40  ft. 
thick,  however,  it  is  only  necessary  to  have  the  panel  pillars  (at  ten-room 
intervals)  of  adjoining  seams  superincumbent,  and  to  lay  out  the  entries 
and  room  axes  of  both  seams  approximately  parallel  to  each  other;  in 
this  way  the  work  in  different  seams  can  be  pursued  more  ndependently 
and  just  as  safely. 

Shaft  Pillars. — The  same  principles  and  formula?  can  be  applied  to 
the  design  of  pillars  for  protection  of  shafts.  In  Fig.  16  the  vertical  shaft 
fd  will  need  a  pillar  in  each  seam  extending  to  the  intersection  with  the 
plane  of  fracture  passing  through  the  shaft  collar  at  /.  Thus  the  mini- 
mum upper  limit  of  these  pillars  must  be  at  e  and  h,  which  for  considerable 
depth,  would  mean  many  hundred  feet  away  from  the  shaft.  But  this 
involves  only  a  moderate  loss  of  ore  because  the  pillar  may  be  narrow 
and  need  extend  only  a  short  distance  down  the  dip  to  b  and  d.  The 
distances  b'b,  d'd  and  the  width  of  the  shaft  pillar  along  the  strike  of  the 
seam  may  be  estimated  by  formula  (4) . 

For  inclined  shafts  following  the  mineral  seam,  the  protecting  pillars 
should  be  continuous  strips  on  each  side  with  break-throughs  only  for  the 
loading  stations.  The  width  of  these  strips,  if  estimated  by  formula  (4), 
should  increase  gradually  from  the  surface  downward.  Although  this 


44 


MINING    WITHOUT    TIMBER 


last  requirement  is  seldom  fulfilled  in  practice,  it  gives  the  minimum  loss 
of  ore  for  the  maintenance  of  a  stable  roof. 

SUPPORT  OF  EXCAVATIONS 

Curved  Sections. — A  tunnel  section  may  be  supported  by  the  circular 
lining  1  (Fig.  17)  against  external  pressure  from  any  direction  since  the 
portion  of  the  ring  taking  the  ground  pressure  will  be  an  arch  to  transmit 
its  load  to  the  balance  of  the  ring  acting  like  arch  abutments.  If  we 
consider  only  the  keeping  open  of  a  given  area  with  the  least  material,  the 
circular  lining  may  be  replaced  with  advantage  by  the  elliptical,  when  the 
pressure  is  greater  in  one  direction  than  in  another,  by  placing  the  long 
axis  of  the  ellipse  parallel  to  the  direction  of  greatest  pressure.  Thus  if 
the  greatest  pressure  comes  from  roof  or  floor,  the  ellipse  should  be  verti- 
cal as  2  in  Fig.  17,  and  if  from  the  sides,  the  ellipse  should  be  horizontal  as 


FIG.  17. — Tunnel  sections. 

in  3.  The  circular  lining  is  most  economical  when  the  external  pressure 
is  equally  distributed,  or  where  it  comes  in  an  oblique  direction,  for  an 
oblique  ellipse  would  be  generally  unsuitable  for  use.  The  oblique  pres- 
sure is  apt  to  occur  when  driving  along  the  strike  of  inclined  beds.  Other 
considerations,  besides  economy  of  lining,  usually  prevail  in  practice  so 
that  circular  sections  are  less  used  than  elliptical  ones,  which  fit  cars 
more  closely  in  transportation  tunnels,  or  egg-shaped  ones,  like  4,  which 
have  a  lesser  hydraulic  gradient  for  water  conduits  or  drains.  The 
material  most  used  for  curved  linings  is  cast  iron  or  some  kind  of  masonry, 
though  steel  shapes  are  also  formed  to  fit,  and  timber  polygons  to  approxi- 
mate curved  sections.  To  merely  support  the  ground,  it  is  clear  that 
only  that  part  of  the  tunnel  section  need  be  lined  which  has  weak  walls 
so  that  we  see  in  practice  linings  on  the  roof  alone,  on  the  roof  and  one 


PRINCIPLES    FOR    CONTROLLING    EXCAVATIONS 


45 


side,  or  on  three  sides,  the  ends  of  the  lining  resting  in  each  case  against 
abutments  of  solid  rock. 

Rectangular  Sections. — The  greatest  available  area  in  transportation 
tunnels  for  the  minimum  volume  of  excavation  is  obtained  by  using  the 
rectangular  instead  of  the  curved  section.  Ordinary  brick  or  stone 
masonry,  having  little  tensile  strength,  is  unsuitable  for  lining  any  part 
of  the  rectangular  section  subject  to  cross-breaking  strains.  Therefore 
it  is  not  used  except  for  side  walls.  Timber  or  steel  beams  and  re-en- 
forced concrete  are  the  common  linings  for  rectangular  sections.  With 
a  weak  roof  and  strong  sides  only  the  piece  ab  (Fig.  18),  which  is  called  a 
cap  or  a  "  quarter-set "  is  put  in;  with  both  roof  and  one  side  weak  the 
cap  ab  and  the  post  be,  called  a  "  half  set,  "  are  needed;  with  roof  and  both 


<r 


FIG.   18. — Tunnel  lining  with  timber. 


sides  weak  a  cap  and  two  side  posts,  or  a  "  three-quarter  set,"  is  used; 
while  with  four  weak  walls  a  cap,  two  posts  and  a  floor  sill,  or  a  "full  set," 
is  required. 

The  attempt  will  not  be  made  here  to  discuss  methods  of  framing  except 
as  they  are  affected  by  ground  pressure.  With  a  predominating  vertical 
pressure  a  good  joint  for  square  timber  is  at  a  (Fig.  18),  and  for  a  round 
timber  at  b.  Where  the  main  pressure  is  horizontal  a  joint  for  square 
timber  is  shown  at  d  and  one  for  round  timber  at  c.  A  rectangular  frame 
can  resist  pressure  acting  parallel  to  its  sides,  but  tends  to  collapse  under 
oblique  pressure.  It  is  to  make  them  more  stable  under  oblique  pressure 
that  tunnel  sets  have  outward-battered  instead  of  vertical  posts. 

Slope  Sections. — In  keeping  open  the  large  stopes  of  some  metal  mines 
with  the  framed  cubical  cells  of  the  square-set  system,  the  same  precau- 
tions of  properly  designed  joints  and  of  uniformly  spaced  points  of  contact 
with  the  surrounding  rock  must  be  observed.  Miners  have  found  to  their 
sorrow  that  it  is  useless  to  attempt  to  keep  open  square-setted  stopes 
under  oblique  pressure  unless  diagonal  braces  (like  bd)  are  inserted,  par- 


46  MINING    WITHOUT    TIMBER 

allel  to  the  direction  of  pressure,  for  transforming  the  unstable  squares 
of  the  frame  into  stable  triangles. 

ZONES  OF  FRACTURE  AND  FLOWAGE 

Wooden  or  steel  frames  will  only  keep  the  peripheral  surface  of 
excavations  intact  against  the  pressure  of  loose  pieces  or  sub-arch  blocks 
like  acb  in  Fig.  13.  For  the  support  of  the  overlying  formation,  even 
masonry  is  only  of  limited  commercial  utility,  therefore  rock  pillars  or 
filling  with  waste  must  be  relied  upon.  Beyond  a  certain  depth,  or  below 
the  "zone  of  fracture"  of  geologists,  we  have  the  "zone  of  flowage," 
where  no  opening  can  be  maintained  permanently  owing  to  the  inability 
of  any  fraction  of  the  rock,  left  as  pillars,  to  sustain  the  superincumbent 
pressure. 

Transposing  formula  (3)  we  have  for  In!  (the  depth  of  the  zone  of 
fracture)  : 


but  for  the  zone  of  flowage  both  m  and  x  are  =  1  and  substituting  these 
values  we  have 

h'  =  --  .............................................   (7) 

w 

From  equation  (7)  it  is  evident  that  the  depth  h'  depends  solely  on 
the  compressive  strength  of  the  basal  rock  and  the  specific  gravity  of  the 
overlying  formation.  Assuming  the  specific  weight  w  of  the  earth's  crust 
to  be  150  Ib.  per  cu.  ft.  and  the  compressive  strength  of  the  basal 
rock  to  be  3,000,000  Ib.  per  sq.  ft.,  we  have  by  substitution  in  (7) 

h,  =3,0(        ^  =  20,000  ft.,  or  about  4  miles. 
150 

In  those  localities  where  the  surface  rock  is  so  porous  as  to  contain  a 
considerable  proportion  of  water,  w  might  be  less  than  150  Ib.  and  the 
depth  of  the  zone  of  fracture  correspondingly  increased. 


CHAPTER  V 
PRINCIPLES  OF  MINE  DRAINAGE 

Those  miners  who  talk  much  of  pumping  but  little  of  drainags 
resemble  those  old-fashioned  doctors  who  spend  all  their  time  on  remediee 
and  neglect  diagnosis.  Instead  of  studying  water  conditions  beforehand 
as  a  basis  for  drainage  equipment,  a  too  common  way  is  to  try  to  fit  the 
pumping  plant  to  the  in-flow  after  it  has  appeared.  This  policy  may 
mean  a  drowned  mine,  and  weeks  of  delay  for  the  installation  of  larger 
pumps  and  the  clearing  from  water;  it  may  mean  a  set  of  makeshift 
pumps  of  the  wrong  size  and  of  low  efficiency  which  may  really  be  wholly 
unnecessary  owing  to  the  feasibility  of  natural  drainage. 

Problems  of  drainage  involve  chiefly  the  four  sciences  of  meteorology, 
geology,  hydraulics  and  mechanics.  From  the  first  we  may  determine 
the  quantity  of  rain  likely  to  fall  on  our  mine  watershed;  from  the 
second  the  conditions  affecting  the  behavior  of  underground  water  in 
the  rocks;  from  the  third  the  laws  governing  the  pressure  and  flow  of 
water;  and  from  the  fourth  the  mechanical  methods  of  unwatering. 

ESTIMATE  OF  WATER  TO  BE  DRAINED 

There  are  multitudinous  mineral  deposits,  each  with  a  special  problem 
of  drainage  of  which  only  some  general  features  will  be  discussed  here 
under  three  cases:  I.  Deposits  in  unconsolidated  rock;  II.  Deposits  in 
stratified  rock;  III.  Deposits  in  massive  rock.  For  any  type  the  water 
encountered  in  mining  operations  will  depend  on  four  factors:  (1)  the 
area  of  contributory  watershed;  (2)  the  moisture  falling  on  watershed; 
(3)  the  moisture  percolating  the  surface  of  watershed;  (4)  the  facilities 
for  underground  water  to  enter  the  mine. 

Case  I.  Deposits  in  Unconsolidated  Rocks. — In  Fig.  19  is  shown  a 
cross-section  of  a  gentle  syndinal  rock  trough  ab  filled  with  alluvium  up 
to  the  surface  cd.  It  is  proposed  to  lower  the  "  water  table  "  or  ground 
water  level  wt  down  to  sump  s  in  order  to  mine  a  placer  deposit  extending 
from  a  to  6.  The  conditions  which  determine  the  quantity  of  water  to  be 
handled  depend  on  two  items;  namely,  the  quantity  of  ground  water, 
and  its  velocity  of  entrance  into  the  workings.  For  the  first  item  we 
have  to  calculate  the  area,  rainfall  and  percolation  of  the  contributory 
watershed,  while  for  the  second  item  the  fact  that  it  will  be  affected  by 
the  method  of  drainage  will  have  to  be  taken  into  consideration.  The 
area  and  rainfall  are  also  the  basis  for  the  calculations  of  the  water-supply 

47 


48  MINING    WITHOUT    TIMBER 

engineer,  but  the  latter  reckons  rather  with  the  run-off  than  with  the 
percolation  which  concerns  the  miner. 

With  underground  conditions  as  represented  in  Fig.  19,  the  area 
of  the  contributory  watershed  evidently  extends  in  width  from  e  to  /, 
and  in  length  from  the  sump  s  to  the  head  of  the  valley,  if  cd  is  a  river 
trough,  or  to  the  bounding  contour  of  the  watershed  if  cd  lies  in  a  lake 
basin.  In  general  the  contributory  watershed  is  all  the  ground  area 
that  drains  toward  the  surface  lying  over  the  sump,  wherever  the  surface 
is  connected  with  the  sump  by  pervious  strata  as  in  Fig.  19.  The  depth 


miil^JJi^^j^] 

N¥^vp^^&g^| 
^s^f^fetS^ 

FIG.   19. — Drainage  in  unconsolidated  rock. 

of  current  rainfall  is  recorded  for  most  localities  in  civilized  countries 
at  the  government  meteorological  stations;  and  in  solving  drainage 
problems,  these  records  should  be  scrutinized  for  the  maximum,  mean 
and  minimum  rainfalls  both  by  months  and  years.  From  this  data,  we 
have  the  rainfall  in  the  wettest  year  or  season  in  contrast  with  that  of 
drouths,  but  it  is  important  also  to  note  what  part  of  the  moisture  falls 
as  snow  and  the  melting  seasons  of  the  latter. 

The  whole  rainfall,  however,  does  not  concern  the  miner.  He  is  con- 
cerned only  with  that  fraction  of  it  which  sinks  into  or  percolates  the 
ground  after  evaporation  and  run-off  have  taken  their  tolls.  Then  if 

area  of  a  watershed  =A  sq.  ft. 

depth  of  moisture  falling  on  a  watershed  —  D  ft. 

volume  of  moisture  falling  on  a  watershed  =  Q  cu.  ft. 

volume  of  moisture  running  off  from  a  watershed  =R  cu.  ft. 

volume  of  moisture  evaporating  from  a  watershed  —E  cu.  ft. 

volume  of  moisture  percolating  a  watershed  =P  cu.  ft. 

fraction  of  moisture  Q  evaporating,  or  evaporation  factor  =  e 

fraction  of  moisture  Q  running  off,  or  run-off  factor  =  r 

we  have, 

but  E  =  eQ  and  R  =  rQ 

so  substitute  in  (1)  and 

Q  =  eQ+rQ  +  P  or  P  =  Q  (I-e-r)          (2) 

but               Q  =  AD  (3) 

hence           P  =  AD  (I-e-r)  (4) 

Evaporation  is  dependent  on  the  state  of  the  atmosphere  and  the 

covering  and  texture  of  the  soil.  The  atmosphere  affects  evaporation 


PRINCIPLES    OF    MINE    DRAINAGE  49 

by  its  changes  in  humidity  and  in  movement.  Both  dryness  and  high 
winds  hasten  evaporation  which  is  usually  compared  for  different  atmos- 
pheres by  observing  water  surfaces.  Thus,  in  the  United  States  the 
mean  annual  evaporation  varies  from  40  in.  in  the  Middle  Atlantic  states 
to  50  in.  on  the  Gulf  of  Mexico,  and  95  in.  at  Yuma,  Arizona. 

In  the  same  locality  the  rate  of  evaporation  which  is  approximately 
equal  for  all  bare  soils*  is  greatly  increased  by  a  cover  of  vegetation. 
Thus  a  5-year  trial  at  Geneva,  New  York,  with  an  average  rainfall  of  23.7 
in.,  gave  its  evaporative  factor  (e  in  Equation  (4))  as  0.64  for  bare 
cultivated  soil,  as  0.71  for  bare  undisturbed  soil,  and  as  0.85  for  sod.  Not 
only  the  heat  of  summer  but  its  vegetation  increases  evaporation,  while 
the  ground  surface  in  winter  acts  much  like  bare  soil  unless  covered  by 
snow  or  ice,  the  daily  evaporation  rate  of  which  in  New  England  is 
.02  in.  and  .06  in.  respectively.  A  less  proportion  of  severe  rains  is 
evaporated  than  of  drizzling  rains,  for  as  a  given  area  has  only  a  limited 
rate  of  evaporation  any  excess  moisture  must  either  run  off  or  percolate. 

The  common  method  of  estimating  the  run-off  is  from  measurements 
of  the  quantity  of  water  flowing  in  the  streams  of  the  watershed.  When 
the  bed  of  a  stream  is  once  mapped  in  section,  a  record  of  its  surface- 
height  readings  renders  possible  a  calculation  of  its  sectional  area  which, 
combined  with  corresponding  readings  of  a  current  meter,  gives  the  data 
for  computation  of  flow.  The  percentage  of  rainfall  found  in  streams, 
evaporation  being  neglected,  depends  both  on  the  slope  of  the  surface 
and  on  its  covering.  For  gently  rolling  land  as  in  Iowa,  the  run-off  factor 
(r  in  Equation  (4))  is  0.33,  for  the  rougher  surface  of  the  Middle  Atlantic 
States  it  is  0.40  to  0.50,  while  in  the  mountain  states  of  Colorado  and 
Montana  it  is  0.60  to  0.70.  The  surface  covering  most  favorable 
to  a  heavy  flow  is  frozen  snow  over  which  over  90  per  cent,  of  the  rainfall 
may  run  into  the  streams,  while  the  melting  of  the  winter's  snow  by  warm 
rains  causes  the  freshets  and  floods  of  spring.  Where  the  surface  is 
irregular  so  that  the  rainfall*  collects  in  ponds  and  swamps  instead  of 
reaching  streams,  the  run-off  is  lessened,  and  the  evaporation  and  per- 
colation is  correspondingly  increased. 

The  beds  of  surface  streams  must  be  relatively  impervious,  for  if  they 
were  freely  percolated  by  water,  there  would  soon  be  no  visible  flow. 
Where  a  stream's  bed  is  is  partially  porous,  much  of  the  water  sinks  to  the 
first  impervious  stratum  and  there  forms  an  invisible  stream  called  the 
underflow  which  often  contains  more  water  than  its  parent  overhead. 
Where  a  stream  has  not  naturally  a  channel  of  impervious  rock  or  clay, 
the  tendency  is  for  it  to  stop  the  pores  of  a  sandy  or  other  pervious  bed 
with  sediment;  especially  is  this  so  in  alluvial  valleys  like  that  of  Fig.  19, 
where  there  might  be  no  visible  stream  at  all  had  the  river  at  r  not  a 
clay-coated  bottom. 

For  our  drainage  problem  of  Fig.  19,  we  have  now  discussed  how  to 


50  MINING    WITHOUT   TIMBER 

ascertain  the  area  of  watershed  A,  the  depth  of  rainfall  D,  the  evapora- 
tive and  run-off  factors  e  and  r,  and  by  substituting  these  values  in  equa- 
tion (4)  we  have  the  percolation  P.  The  result  from  solving  Equation 
(4)  can  be  compared  with  the  following  table  which  gives  for  the  percen- 
tage of  total  rainfall  percolating  various  surfaces : 

sand  =  60  to  70 
chalk  or  gravelly  loam  =  35 
sandstone  =  25 
limestone  =  15 
clay  or  granite  =  15  and  "ess 

We  do  not  have  to  provide  at  s  for  the  drainage  of  volume  P,  but  only 
for  that  portion  of  it  which  is  not  drained  off  elsewhere,  does  not  reascend 
to  evaporate  at  the  surface  or  is  not  held  in  the  pores  of  the  subsoil. 
The  drainage  elsewhere  would  be  nil  in  a  lake  basin  with  impervious 
bottom,  but  in  the  usual  self-draining  basin  it  would  constantly  tend  to 
lower  the  water  table. 

The  lake-basin  condition  is  well  exemplified  at  both  Bisbee  and 
Tombstone,  Arizona.  These  camps  lying  in  the  Mule  Mountains,  where 
the  annual  rainfall  is  under  12  in.  and  the  evaporative  factor  large,  would 
be  casually  reckoned  as  having  dry  mines,  but  the  very  opposite  is  the 
case.  The  ore  bodies  in  each  camp  are  found  in  limestone  and  shale 
beds  which  are  so  folded  as  to  form,  with  the  adjoining  intrusive  rocks, 
an  impervious  basin  which  catches  all  the  rain  percolating  the  surface  over 
a  large  watershed.  The  present  water  in  the  basins  represents  the 
accumulation  of  years,  so  to  lower  it  has  taken  more  pumping  than 
would  be  necessary  in  a  very  wet  valley  whose  ground  water  was  depend- 
ent solely  on  current  rainfall.  At  Bisbee  in  1906  the  water  level  had 
been  permanently  lowered,  for  three  years'  pumping  by  several  companies 
had  reduced  the  inflow  at  the  Calumet  and  Pittsburg  shaft  from 
3000  to  1500  gal.  per  min.;  but  at  Tombstone  in  1911,  where  nine  years 
of  pumping  of  the  Contention  shaft  had  little  affected  the  original  flow 
of  3000  gal.  per  min.,  it  was  deemed  unprofitable  to  struggle  further  and 
the  pumps  were  pulled.  The  excess  of  water  in  the  Tombstone  basin 
probably  comes  from  an  adjoining  watershed  through  underground 
channels. 

The  loss  of  ground  water  by  evaporation  increases  with  a  damp  soil 
and  a  high  water  table.  Consequently,  in  self-draining  ground  the 
evaporation  is  greater  if  the  rainfall  is  evenly  rather  than  sporadically 
distributed.  Evaporation,  however,  usually  affects  the  water  table  only 
slightly  as  compared  with  the  capacity  of  the  formation  for  the  storage, 
surrender  and  passage  of  water. 

In  Fig.  19,  the  section  of  the  water-storage  area  between  wt  and  bed- 
rock is  not  the  whole  area  wmnt,  but  this  area  multiplied  by  the  factor 


PRINCIPLES    OF    MINE    DRAINAGE  51 

for  " voids"  or  the  proportion  of  intergranular  spaces  in  the  formation. 
The  void  factor  depends  less  on  the  size  of  rock  grains  than  on  their  uni- 
formity, and  varies  from  0.2  to  0.5.  Yet  another  item  must  be  included, 
in  estimating  the  quantity  of  water  that  must  be  drained  to  lower  wt,  and 
that  is  the  factor  for  surrender  or  yield  which  depends  on  the  capillarity 
or  fineness  of  the  grain  of  the  formation.  The  yield  factor  is  almost  nil 
for  clay,  0.5  to  0.6  for  porous  soil,  0.6  to  0.7  for  sand,  and  nearly  1.0  for 
clean  gravel  or  boulders.  A  low  yield  factor  means  not  only  the  reten- 
tion of  rainfall  in  a  porous  formation  until  it  is  saturated,  but  a  long  delay 
before  a  heavy  shower  begins  to  be  noticed  underground.  From  the 
above,  if 

W  =  volume  in  cu.  ft.   of  water-bearing  formation  tributary  to 

sump  s 
W  =  volume  in  cu.  ft.  of  water  the  water-bearing  formation  yields 

tributary  to  sump  s 
x  =  factor  for  voids  in  formation 
y  =  factor  for  water-yield  of  formation 

then  W  =  xyW.  (5) 

To  free  our  placer  ab  from  water,  it  will  not  be  necessary  to  lower  the 
water  table  to  the  profile  wmnt,  but  only  to  the  profile  whabgt  where  who, 
and  bgt  are  the  profiles  of  the  hydraulic  gradient  toward  the  sump  s.  The 
hydraulic  gradient  increases  with  the  fineness  of  grain,  though  very  small 
in  gravel,  it  is  30  to  50  ft.  per  mile  in  sand  and  in  a  large  basin,  it  would 
thus  considerably  decrease  the  volume  of  water  tributary  to  sump  s. 

The  hydraulic  gradient  for  a  given  formation  can  be  directly  measured 
by  digging  two  wells  in  the  same  line  of  water  drainage,  at  some  distance 
apart,  and  then  recording  their  water  levels.  The  hydraulic  gradient 
vvill  then  be  the  difference  of  water  level  divided  by  the  distance  between 
the  wells. 

Hajzen  gives  as  a  formula  for  the  velocity  of  passage  of  ground  water 
V  =  KD2S  (6) 

where  V  =  velocity  in  ft.  per  sec.  of  flow  through  ground  pores 
#  =  0.29  (a  constant) 

D  =  diameter  in  mm.  of  sand  grain  "  effective  "  (i.e.,  90  per  cent,  of 
grains  must  be  larger  than  D) .  Formula  (6)  is  inapplicable 
when  d  is  less  than  3 

£  =  sine  of  slope  of  the  hydraulic  gradient 

Then  if  B  =  area  in  sq.  ft.  of  a  vertical  surface  enclosing  mine 
openings  extending  from  water  surface  in  pump  to  water 
table.  Height  of  surface  B  should  be  small  for  use  of 
formula  (6) 

/=  volume  of  water  in  cu.  ft.  entering  placer  per  sq.  ft.  of  area  B 
F  =  total  volume  in  cu.  ft.  entering  placer  over  total  area  B 
kf  =  fractional  factor  for  voids  in  walls  of  area  B 


52  MINING    WITHOUT   TIMBER 

It  is  evident  that 

f=k'V  andF  =  k'BV 
Substitute  for  V  from  Equation  (6)  and 

F  =  kK'BD2S  (7) 

In  practice  the  possibility  of  keeping  the  placer  dry  enough  to  permit 
miners  to  work  would  of  course  depend  not  only  on  the  means  of  drainage 
available  to  keep  sump  s  clear,  but  also  on  /  or  the  rate  of  inflow  at  the 
mining  face.  As  /  increased  beyond  a  certain  figure,  the  miners  would 
find  themselves  working  in  a  heavy  spray  and  standing  in  a  gurgling  pond. 
In  such  a  case,  unless  the  inflow  could  be  controlled  by  a  cofferdam,  sub- 
aqueous mining  would  have  to  be  resorted  to. 

Case  II.  Deposits  in  Stratified  Rock. — An  example  of  this  case  is 
shown  in  Fig.  20,  a  cross-section  of  a  coal  seam  A  in  a  synclinal  basin. 
Beneath  the  coal  is  a  thin  layer  of  clay  B  resting  on  a  sandstone  C,  and 


FIG.  20. — Drainage  in  stratified  rock. 

above  it  are  strata  of  sandstone  D,  shale  G,  and  limestone  H.  Along  the 
surface  runs  a  river  r  over  a  valley-filling  of  alluvial  soil.  Then  the  perco- 
lation into  a  coal  mine  at  A  will  depend  not  only  on  the  coal  itself  whose 
bedding  and  joint  planes  may  be  somewhat  permeable,  but  on  the  nature 
of  the  adjoining  rocks. 

Clay  and  shale  are  not  only  relatively  impermeable  but  plastic,  and 
tend  to  close  tm  any  openings  made  inyaving  aorogenic  movements. 
Sandstones  vary  in  their  structure,  some  h  duebh  texture  as  porous  as 
free  sand,  while  the  grains  of  others  are  closely  cemented  and  almost 
impermeable.  Limestones,  especially  if  dolomitic,  abound  in  irregular 
channels  and  pot-holes,  often  large  enough  to  contain  underground  rivers 
or  ponds.  Should  the  rocks  of  Fig.  20  be  subjected  to  metamorphism, 
their  permeability  would  be  much  diminished,  or  perhaps  entirely  de- 
stroyed, as  pores  and  bedding  planes  were  obscured,  until  we  approached 
as  the  limit  the  massive  formation  of  Case  III.  Clay  and  shale,  when 
metamorphosed,  become  dense  and  strong  slate  or  schist,  sandstone  solidi- 
fies into  impermeable  quartzite,  and  limestone  changes  into  crystalline 
marble. 

From  these  considerations,  it  can  be  seen  that  the  stratum  of  shale 
at  G,  provided  it  has  not  been  pierced  by  orogenic  movements  or  human 
hand,  acts  as  a  screen  to  keep  out  any  water  which  may  percolate  into 
the  limestone  from  the  watershed  ef.  As  the  strata  outcrop,  however, 


PRINCIPLES    OF   MINE    DRAINAGE  53 

beyond  the  summits  e  and  /  of  the  synclinal  basin,  the  coal  seam  will  be 
exposed  to  percolation  from  watersheds  ec  and  fc'. 

As  long  as  the  impermeable  clay  floor  B  of  the  coal  is  uncracked,  the 
watersheds  contributory  to  the  coal  seam  will  extend  only  from  e  to  6  and 
from  /  to  61  and  not  all  of  their  percolation  will  reach  the  coal,  because 
the  shale  stratum  G  will  seal  off  any  surface  water  that  may  enter  the 
limestone  layer  H  between  the  crests  e  and  /  and  tire  roof  of  G.  Should 
the  floor  B  be  cracked  or  feathered  out  in  places,  it  may  be  serious  from 
a  drainage  standpoint,  for  the  coal  seam  will  then  be  open  to  a  flood  from 
the  hydrostatic  water  in  sandstone  C.  Thus  in  the  rock  formation  of 
Fig.  20,  the  ground  water  would  not  occur  in  a  connected  body  as  in  Fig. 
19.  but  each  porous  zone  would  contain  its  own  pool  separated  from  the 
others  by  an  impermeable  stratum.  The  equivalent  of  the  water  table 
wt  of  Fig.  19  would  be  found  here  in  the  limestone  H,  but  it  would  cir- 
culate there  in  irregular  open  channels  instead  of  in  intergranular  pores. 
The  contributory  watersheds  having  been  thus  measured,  we  have  only 
then  to  gather  the  other  meteorological  and  physical  constants,  as  ex- 
plained for  Case  I,  in  order  to  solve  Equations  (1)  to  (7)  for  the  drainage 
of  Fig.  20. 

Case  III.  Deposits  in  Massive  Rock. — In  Fig.  21,  let  cd  be  the  cross- 
section  through  a  fissure  vein  in  massive  rock,  which  is  either  of  igneous 


FIG.  21. — Drainage  in  massive  rock. 

origin  or  so  metamorphosed  that  its  sedimentary  pores  and  bedding  planes 
are  practically  obliterated.  This  formation,  then,  instead  of  being  quite 
'porous  like  that  of  Case  I  or  irregularly  porous  like  that  of  Case  II,  is  in 
its  original  condition,  more  or  less  impermeable,  but  in  mining  regions 
it  has  usually  been  so  cracked  by  earth  movements  as  to  abound  in  open- 
ings which  grade  from  wide  fissures,  both  long  and  deep,  to  such  minute 
fracture  planes  as  those  of  the  Bingham  copper  porphyry  which  scarcely 
pass  seepage  water.  When  rainfall  can  only  percolate  the  surface  of 
Fig.  21,  through  irregularly  spaced  crevices  or  joints  instead  of  through 
a  porous  zone,  there  can  be  nothing  like  a  general  ground  water  level 
except  within  areas  whose  crevices  are  all  connected.  Thus  each  crevice 
system  has  a  height  of  water  table  varying  according  to  the  size  and  nature 


54  MINING    WITHOUT    TIMBER 

of  its  contributory  watershed.  The  mineral  veins  themselves  have  often 
trunk  channels  along  their  walls  which  receive  water  from  numerous 
branch  cracks  and  fissures. 

The  watershed  tributary  to  vein  cd  of  Fig.  21  will  not  extend  laterally 
from  e  to  /  as  in  Case  I  unless  all  the  intermediate  fissure  systems  lead  to 
the  vein,  but  it  may  cover  a  much  wider  area  owing  to  the  possible 
juncture  of  subterranean  streams  with  cd,  which  streams  in  mountainous 
regions  may  be  under  a  high  hydrostatic  head.  In  fact,  only  the  map- 
ping of  the  region's  underground  water  channels,  and  this  could  seldom  be 
done  except  in  an  extensively  developed  district,  would  enable  an  engi- 
neer to  satisfactorily  solve  Equations  (1)  to  (7)  as  in  the  two  previous 
cases.  In  mining  cd,  care  would  have  to  be  taken  on  the  hanging  side, 
for  by  the  tapping  of  natural  blind  crevices  or  by  allowing  the  hangwall 
to  move  and  crack,  the  river  r  might  be  precipitated  into  the  workings. 

It  is  probable  that  some  of  the  hot  water  found  in  mining  such  igneous 
formations  as  the  Comstock  lode  comes,  not  from  rainfall,  but  directly 
from  the  occluded  moisture  of  cooling  magmas.  According  to  the  nebu- 
lar hypothesis,  all  surface  water  had  originally  an  igneous  origin.  The 
miner  who  operates  in  a  region  of  magmatic  water  cannot  estimate  its 
quantity  beforehand,  as  in  the  case  of  meteoric  inflows,  but  must  simply 
handle  it  as  it  appears. 

As  mines  get  deeper  and  rock  pressures  become  greater,  fissures  and 
other  open  spaces  tend  to  close  up  and  long  before  the  bottom  of  the  zone 
of  crust  fracture  is  reached,  at  a  depth  of  something  around  4  miles,  there 
is  little  free  water  in  the  rocks.  At  the  Calumet  and  Hecla  Copper  mine, 
Mich.,  in  a  conglomerate  lode  bedded  between  amygdaloids,  the  maxi- 
mum water  flow  is  at  1800  ft.  along  the  38  deg.  dip,  while  at  3000  ft.  the 
water  flow  is  insufficient  even  to  supply  the  drills. 

CONTROL  OF  WATER 

This  topic  naturally  divides  itself  into  surface  and  underground 
control. 

Surface  Diversion. — It  is  much  better  to  keep  water  out  of  a  mine  than  to 
use  the  most  approved  method  of  drainage  after  its  unnecessary  entrance^ 
Surface  run-off  is  kept  out  of  a  mine  ditching  around  shafts  and  vein  out- 
crops as  C  in  Fig.  21.  Often  it  is  best  to  refrain  from  stoping  a  vein  out 
quite  up  to  the  surface  in  order  to  keep  rain  out  of  the  workings.  A 
stream  above  the  mine,  which  seeps  badly  into  its  bed  or  whose  bottom 
may  be  cracked  by  caving  operations,  can  often  be  diverted  to  another 
channel  or  carried  in  a  flume  over  the  dangerous  stretch. 

Underground  Diversion. — In  Fig.  20  the  penetration  of  the  imperme- 
able shale  layer  G  by  the  shaft  A  A'  will  eventually  drain  the  wet  lime- 
stone layer  H  into  the  mine  at  A  unless  some  precaution  is  taken.  Two 


PRINCIPLES    OF    MINE    DRAINAGE  55 

remedies  suggest  themselves,  the  first,  a  concrete  shaft-lining  from  the 
surface  down  to  a  sealed  footing  in  the  shale;  the  second,  the  usual  pervi- 
ous shaft-lining  provided  with  a  water  ditch  or  "ring"  around  the  shaft, 
in  the  roof  of  the  shale,  which  catches  the  water  from  above  and  leads  it 
to  a  sump,  which  has  means  for  drainage,  on  the  same  level. 

In  ore  deposits  in  hilly  regions,  an  impervious  floor  sometimes  has 
below  it  a  sandstone  or  other  porous  stratum  which  dips  toward  an  out- 
crop, on  a  hillside  at  a  lower  level,  and  is  thus  self-draining.  In  such  a 
case,  diamond-drill  holes  or  a  winze  through  the  floor  of  the  mine  sump 
into  the  porous  stratum  will  effectively  drain  the  workings. 

Natural  Dams. — Rock  barriers  are  highly  useful  in  the  control  of 
water  in  mines.  As  already  explained,  where  tight  strata  cut  off  the 
mine  from  wet  formations,  such  natural  seals  should  be  left  undisturbed 
if  possible.  Pillars  of  mineral  are  often  left  between  adjoining  mines  to 
keep  their  water  systems  distinct,  and  in  many  states  a  barrier  about 
50  ft.  wide  must  be  left  unmined  around  the  boundary  of  coal  properties. 

The  Lehigh  Valley  Coal  Co.  is  now  mining,  near  Hazleton,  Pa.,  a 
synclinal  trough  containing  parallel  anthracite  seams  which  extends  for 
several  miles  and  dips  for  about  3000  ft.  vertically  in  that  distance.  The 
trough  has  been  divided  into  three  drainage  basins  by  leaving  a  trans- 
versal barrier  pillar  of  coal,  100  ft.  wide,  below  each.  The  barriers  are  at 
altitudes  of  1084,  1250,  and  4000  ft.  respectively  and  each  has  its  own 
unwatering  system.  Each  barrier  is  pierced  by 
boreholes  lined  with  pipes  whose  valves  can  be 
opened  to  drain  the  basin  above  into  the  one  below 
in  case  of  an  emergency. 

It  is  often  necessary  to  penetrate  water  barriers 
in  order  to  drain  old  mines,  and  where  the  dammed- 
up  water  is  under  a  high  head  it  is  best  tapped  by 
drilling.  Boring  long  holes  for  tapping  can  be 
done "  in  any  direction  by  a  diamond  drill.  A 

J  „  .  .  FIG.    22.— W,- 

customary  safe-guard  against  heavy  pressure  is  to     iron  Mt.  Mme, 
bore  the  first  few  feet  of  the  hole  large  enough  for 

a  pipe  lining  ck,  Fig.  22,  whose  exterior  is  made  to  fit  the  rock  tightly 
by  a  packing  of  lamp-wick,  wound  spirally,  or  of  cement.  When  it  is 
necessary  to  regulate  the  flow  from  the  hole,  a  valve  v  is  put  on  the 
lining  pipe  whose  end  must  then  be  anchored  to  posts  p.  Such  a  valve 
on  a  drill-hole  flow,  small  anyhow,  allows  the  cautious  emptying  of  old 
workings  where  a  sudden  release  of  water  might  damage  the  shaft  or 
other  important  pillars. 

For  short  distances,  tapping  can  be  well  .performed  with  a  percussive 
drill  and  a  typical  recent  case  can  be  cited  at  the  Iron  Mt.  Mine,  Montana, 
where  the  new  drainage  adit  was  connected  with  the  old  shaft-workings 
which  contained  some  100,000,000  gal.  of  water  under  a  head  of  900  ft. 


56 


MINING    WITHOUT    TIMBER 


When  the  face  of  the  tunnel  ta,  Fig.  22,  arrived  near  the  old  shaft  s,  a 
6x7-ft.  cross-cut  tc  was  run  for  30  ft.,  parallel  to  the  shaft  station,  and 
from  c  a  taildrift  was  carried  back  for  30  ft.  to  d.  Next  a  3-in.  percussive 
drill  was  set  up  at  c  and  a  hole  drilled  in  for  10  ft.  to  admit  a  4-in.  pipe 
lining  ck  which  was  then  well  cemented  and  anchored  in.  Drilling  was 
proceeded  beyond  the  lining  with  a  1  3 /4-in.  bit  when,  at  e,  23  ft.  from 
the  collar,  the  point  holed  through.  On  loosening  the  chuck,  the  bit 
was  shot  back  by  the  water  pressure  against  the  end  c,  and  was  followed 
by  a  swift  stream  of  water,  but  as  a  low  dam  had  been  erected  across  the 
crosscut  at  t,  the  men  climbed  over  it  and  safely  reached  the  adit  mouth, 
a  mile  and  a  half  distant. 

Artificial  Dams. — Many  dams  are  built  underground:  for  making 
sumps  out  of  old  headings  or  stopes;  for  regulating  the  flow  to  the  pumps; 
for  isolating  the  water  of  abandoned  workings;  and  for  confining  water  to 


Section  x-y 
FIG.  23. — Artificial  dams. 

certain  localities,  as  in  the  case  of  flooding  mine  fires  or  of  filling  seams 
by  the  flushing  system.  Mine  dams  differ  from  those  on  the  surface  in 
the  fact  that  they  often  stop  opsnings  of  small  height  relative  to  the 
pressure  of  water  to  be  sustained.  In  such  cases,  mine  dams  must  have  a 
solid  footing  all  around  their  periphery  instead  of  just  at  the  base  and 
sides  like  a  river  dam.  Favorite  materials  for  dams  are  wood,  brick, 
stone  and  concrete. 

A  diverting  dam  whose  crest  is  higher  than  the  water  surface  it  sustains 
can  be  built  light  and  like  a  surface  structure;  but  precautions  must  be 
taken  to  successfully  sustain  a  high  water-head  (which  causes  a  pressure  of 
0.434  Ib.  per  sq.  in.  for  each  foot  of  height),  and  the  arch  is  a  favorite  form 
for  this  purpose.  Fig.  23  shows  a  composite  plan  and  section  of  a 
dam,  to  back  up  water  at  y  across  a  heading  or  shaft,  which  is  made  of 
two  arches,  ab  and  cd,  with  a  filling  between  of  puddled  clay  or  concrete. 
It  will  be  noticed  that  the  heading  walls  are  cut  out  to  give  indented  skew- 
backs  for  the  arches  except  at  e'/  and  g'  h'  where  a  plastic  roof  and  floor 


PRINCIPLES    OF    MINE    DRAINAGE.  57 

might  make  indentation  unnecessary  if  the  swelling  wood  construction, 
to  be  described  later,  were  used. 

Both  drain-pipe  at  m  and  air-escape  at  n  are  provided  with  valves  and 
sealed  tightly  in  the  structure.  The  manhole  pipe  xy  is  anchored  to  the 
dam  and  is  as  essential  during  construction  as  afterward.  With 
moderate  pressure  one  arch,  like  eh,  is  enough;  and  to  build  it  of  wood, 
each  piece  should  be  the  length  of  eh  and  tapered  wedge-shape  like  an 
arch  stone.  A  tight  joint  can  be  made  between  wood  and  walls  by 
tarred  felt  and  small  wedges,  and  the  pipes  can  be  sealed  in  with  wedges. 
When  of  masonry,  the  arches  are  laid  over  wooden  centers,  the  under  one 
of  which  is  left  in  permanently  if  the  dam  is  across  a  shaft.  Masonry 
dams  are  kept  tight  by  a  concrete  or  clay  backing,  and  as  the  latter 
needs  to  be  confined  under  heavy  pressure,  the  double  arches  of  Fig.  23, 
with  clay  between,  are  then  especially  suitable.  Flat  wooden  dams 
are  often  used  and  usually  they  are  held  by  posts  with  ends  hitched  into 
the  walls.  The  wooden  lining  is  made  of  several  layers  of  planks  and, 
with  walls  too  soft  for  its  support  by  posts  at  intervals  in  hitches,  the 
lining  itself  may  be  extended  into  brick-lined  hitches  cut  in  the  walls, 
and  its  central  portion  be  backed  by  a  timber  set  whose  battered  posts  are 
set  in  the  direction  of  pressure  and  rest  in  hitches  cut  in  the  heading's 
walls. 

At  the  Chapin  iron  mine,  Michigan,  dams  have  been  helpful  in  the  con- 
trol of  a  big  inflow  of  water.  The  Chapin  ore-body  is  a  hematite  lense 
appearing  in  cross-section  about  like  cd  in  Fig.  21.  It  is  enclosed  by  slate 
walls  but  has  an  extensive  dolomite  formation  about  100  yards'  distant  on 
the  hanging  side.  The  author  found  on  his  visit  in  1908  that  the  flow  at 
the  1000 -ft.  level  had  not  been  appreciably  lessened  in  spite  of  pumping 
2000  to  3000  gal.  per  min.  for  the  previous  seven  years.  The  water 
proceeds  from  channels,  in  the  dolomite  hangwall,  which  are  believed  to 
connect  with  two  small  lakes  several  miles  to  the  northeast.  If  the 
originally  impermeable  slate  hangwall,  that  cut  off  the  ore  from  the 
water-bearing  dolomite,  had  not  been  cracked  for  over  300  ft.  from  the 
surface  (by  the  caving  operations),  the  drainage  problem  could  easily 
have  been  solved  by  keeping  shafts  and  cross-cuts  entirely  in  foot-wall. 

The  No.  2  Hamilton,  vertical  shaft,  then  used  for  pumping,  had  been 
sunk  in  the  hanging  dolomite  and  great  difficulty  had  been  encountered 
in  driving  the  1000-ft.  and  lower  cross-cuts  because  of  the  water  crevices 
encountered.  In  starting  the  1000-ft.  cross-cut  a  compound  station  pump 
was  first  installed;  but,  nevertheless, the  first  water  crevice  struck  had  to 
be  dammed  with  masonry,  and  the  pressure  gauge  showed  a  static  head 
there  from  a  water  level  within  300  ft.  of  th3  surface.  Next,  a  branch 
drift,  some  distance  back  from  dam  No.  1,  was  begun;  but  this  also 
struck  a  crevice  and  had  to  be  dimmed. 

A  second  branch  drift  was  then  started  and  dammed  (after  only  a 


58  MINING    WITHOUT    TIMBER 

short  advance)  and  a  second  compound  pump  installed  at  the  station. 
This  last  dam  was  fitted  with  a  water-tight  iron  door  opening  outward, 
so  when  drifting  was  continued  beyond  it  (to  make  a  chamber  for  diamond 
drilling)  the  excavated  earth  could  be  passed  back  in  boxes.  With  the 
diamond  drill,  the  space  yet  to  transverse  to  reach  the  vein  was  searched 
for  a  cross-cut  opening  free  from  crevices;  but  as  none  was  found  the 
cross-cut  had  finally  to  be  finished  anyhow  by  the  aid  of  strenuous 
pumping. 

DAMMING  BY  DEPOSITION 

E.  B.  Kiri>y  has  devised  a  method  (U.  S.  Pat.  No.  900,683)  of  sealing 
the  rock  crevices  of  mine  workings  by  the  deposition  therein  of  sedi- 
ment. Finely  divided  clay  is  preferable  but  other  materials  may  be 
used  such  as  sand,  mill  tailing  or  slime,  cement,  saw-dust,  horse  manure, 
chopped  hay  or  fiber.  The  injection  of  the  water  bearing  this  material 
in  suspension  may  be  made  by  force-pumps,  or  by  stand-pipes  extending 
far  enough  toward  the  surface  to  furnish  the  necessary  pressure. 

The  suspended  particles,  when  put  in  a  cavity  containing  water  in 
motion  toward  exits  in  the  mine,  are  seized  and  carried  toward  such 
exits,  settling,  accumulating  in,  and  choking  at  various  points  the  con- 
tributory passages.  The  moving  currents  automatically  select  those 
passages  which  are  discharging  water  into  the  mine  and  require  sealing; 
they  disregard  other  passages  because  the  water  is  not  in  motion  in 
them.  The  choking  which  occurs  in  the  outflowing  passages  is  gradual 
and  at  those  most  favorable  points  where  the  passages  are  smallest  and 
the  flow  most  diffused.  In  fact  large  passages  cannot  be  thus  choked 
but  must  by  dammed. 

When  the  flowing  passages  are  choked  the  process  ceases  even  though 
other  passages  are  still  open.  If  by  the  choking  of  one  or  more  passages 
the  current  is  deflected  to  others,  the  deposition  is  there  automatically 
resumed.  At  any  choked  locality  the  water  pressure  holds  the  choking 
particles  firmly  in  place  and  produces  a  perfect  seal  by  shutting  off  the 
threads  of  water  in  every  contributing  passage. 

Adits. — These  are  tunnels  run  in  from  a  low  surface  point  to  drain 
underground  workings.  In  Fig.  21,  it  is  evident  that  the  adit  ad  would 
drain  all  of  vein  cd  above  level  d  and  that  adit  bk  would  drain  everything 
above  level  k.  The  water-bearing  fissure  mn  cuts  the  vein  at  n,  and  to 
drive  an  adit  at  the  level  n  would  obviously  be  impractical  with  the  given 
topography;  but  nothing  would  hinder  the  extension  of  adit  bk  to  the 
fissure,  and  the  running  of  a  drift  g  along  its  footwall,  as  far  as  necessary, 
to  intercept  all  the  water  drainage  into  the  mine  at  n.  This  scheme  was 
employed  to  supplement  the  lower  adit  of  the  Horn  Silver  mine  in  Utah. 

The  following  remarks  apply  to  all  drainways,  whether  adits  proper, 
debouching  at  the  surface,  or  merely  interior  tunnels  emptying  into  a 


PRINCIPLES    OF    MINE    DRAINAGE 


59 


sump.     The  minimum  grade  for  long  modern  adits  with  unlined  ditches 
is  1/4  of  1  per  cent.     The  carrying  power  of  water  channels  can  be  thus 
estimated: 
If 

Sine  of  slope  of  hydraulic  gradient  of  water  flowing  in 

channel 

Area  of  -cross-section  of  water  flowing  in  channel 
Velocity  per  sec.  of  water  flowing  in  channel 
Wetted  perimeter  of  the  containing  surface  of  channel 
Constant,  increasing  with  smoothness  of  containing  sur- 

face of  channel 

Quantity  of  water  flowing  per  sec.  in  channel 
then  from  Merriman's  "  Hydraulics  " 

loS 


S 

a  sq.  ft. 
v  ft. 
P  ft. 

c 

q  cu.  ft. 


(8) 


but    =  av 


laS 
hence  q=ae\\— 


(9) 


In  those  cases  where  adits  are  only  to  be  used  for  drainage,  a  circular 
section  is  often  preferable;  because  it  carries,  when  running  half  full  or 
more,  the  most  water  for  a  given  volume  of  exca- 
vation; is  stable  against  external  pressure;  and  is 
readily  adaptable  to  masonry  lining,  which,  be'ng 
smoother  than  wooden  sets,  gives  a  larger  value  for 
c  in  formula  (9),  and  consequently  passes  more 
water.  Where  the  water  deposits  sediment,  the 
egg-shaped  section  of  Fig.  17  (4),  used  for  sewers, 
best  enables  a  uniform  carrying  power  to  be  main- 
tained as  the  height  of  water  fluctuates. 

-  When  adits  serve  for  haulage  as  well  as  drain- 
age, the  economical  shape  is  usually  oval  or  rec- 
tangular. The  oval  shape  is  best  for  weaker  walls 
with  external  pressure  mainly  vertical,  and  it  can 
easily  be  lined,  where  necessary,  by  masonry.  The 
rectangular  shape  is  common  where  the  adit  fol- 
lows some  flat  stratum  like  a  coal  seam  with  a  strong  roof,  that  will  stand 
without  arching;  or  where  most  of  the  length  has  to  be  supported  by 
timber  or  metal  sets  which  are  ill  adapted  to  curved  sections. 

A  compromise  section  for  an  adit  is  shown  in  Fig.  24  (a)  with  a  self- 
sustaining  arched  roof  and  a  flat  bottom  to  give  a  cheap  footing  for  the 
track  ties.  With  a  moderate  amount' of  water  it  can  be  carried  in  a  side 
ditch  which  is  easier  to  watch  and  clean  than  one  under  the  track.  Where 
the  adit  of  Fig.  24  (a)  is  in  a  narrow  vein  of  width  from/'  toe?',  the  ditch 


24.  —  Adit  sections. 


60  MINING    WITHOUT   TIMBER 

is  best  placed  along  that  side  whose  cutting-out,  to  give  space  for  the 
adit,  will  admit  the  least  water  from  the  walls.  The  tightness  of  the 
ditch's  bottom  rock  against  seepage  should  also  be  considered,  if 
there  are  to  be  workings  underneath  it,  and  sometimes  a  wooden  or 
concrete  lining  may  be  necessary  as  commonly  it  is  for  sumps.  Where 
the  adit  can  be  placed  between  vein  walls  as  ef  and  ed  without  cutting 
them,  it  is  usually  best  to  have  the  main  ditch  along  the  footwall  at  c; 
and  connect  it,  if  necessary,  by  cross  ditches  to  an  auxiliary  ditch  along 
the  hangwall  at  ef. 

Both  ditches  and  swamps  should  be  covered  in  hot  mines  like  those 
of  the  Comstock  lode  in  order  to  prevent  any  unnecessary  humidifying 
of  the  air. 

The  new  Roosevelt  adit  at  Cripple  Creek,  Colo.,  will  be  over  3  miles 
long  and  used  only  for  drainage.  This  gold  mining  district  lies  in  an 
igneous  formation,  and  as  it  occupies  an  area  of  about  8  sq.  miles,  it  is 
estimated  that  each  foot  in  height  of  its  ground  water  means  60,000,000 
gal.  of  water.  The  adit  was  started  with  a  section  like  Fig.  24  (a),  10  ft. 
high  and  6  ft.  wide,  but  it  has-been  changed  to  one  6  ft.  high  by  10  ft.  wide 
to  give  space  for  a  curved  ditch,  6  ft.  wide  by  3  ft.  deep,  and  a  narrow 
track  along  one  wall. 

Where  side  ditches  are  inconvenient  or  inadequate,  they  can  be  re- 
placed or  supplemented  by  a  central  ditch  nc'  (shown  dotted  in  Fig.  24  (a)) 
cut  under  the  rails.  For  heavy  flows,  however,  the  whole  bottom  of  the 
adit  may  be  utilized.  In  that  case,  it  should  be  cut  round,  as  ghh'k  in 
Fig.  24  (b),  in  order  to  obtain  the  cheapest  rock  breaking  and  the  maxi- 
mum carrying  power  for  a  given  sectional  area;  unless  a  flat-bedded  forma- 
tion makes  the  excavation  of  the  larger  square  are  a  gg'k  'k  just  as  economical. 
The  track  ties  may  be  spiked  to  stringers  which  are  set  on  posts  or  brick 
piers,  h  and  h't  of  sufficient  height  to  keep  the  ties  above  the  high-water 
flow.  In  double  track  adits,  three  rows  of  stringers  on  piers  are  sufficient 
if  long  ties  are  used.  Where  the  track  is  far  above  the  rock  bottom  and 
the  adit  is  narrow,  cross  beams  like  gk,  hitched  into  the  walls,  may  be  the 
cheapest  supports  for  the  stringers. 

Adits  are  especially  advantageous  in  mountainous  regions  of  steep 
slopes,  where  a  great  height  can  be  drained  with  a  short  adit.  The  only 
drainage  expense  with  adits  is  for  interest  and  maintenance,  and  if  well 
constructed,  they  are  not  subject  to  the  breakdowns  of  mechanical  appa- 
ratus at  critical  moments.  When  the  adit  mouth  is  some  distance  higher 
than  the  stream  into  which  it  drains,  the  escap  ng  water  can  be  effectively 
utilized  for  power.  In  a  wet  district  of  large  producing  mines  whose 
drainways  can  easily  be  connected,  it  is  often  advisable  to  drive  a  very 
long  adit  for  general  drainage. 

Notable  among  such  modern  American  adits  are,  in  Colorado,  the 
Roosevelt  and  the  5-mile  Newhouse  at  Idaho  Springs;  and  in  the  an- 


PRINCIPLES    OF    MINE    DRAINAGE  61 

thracite  region  of  Pennsylvania,  the  5-mile  Jeddo-Basin  in  Luzerne  Co., 
the  1-mile  Oneida  in  Schuylkill  Co.,  and  the  1  1/2-mile  Lausanne  near 
Mauch  Chunk.  The  last  named  drains  13  miles  of  underground  tunnels 
and  14  different  collieries. 

Siphons. — In  mining  flat  coal  and  other  seams,  convex  rolls  often 
occur  in  the  floor  of  the  gangways  which  dam  up  the  drainage.  It  is 
feasible  to  pass  a  low  roll  by  deepening  the  water  ditch;  but  a  high  roll, 
unless  it  is  advantageous  to  also  cut  the  whole  gangway  through  it  to 
obtain  a  uniform  haulage  grade,  is  often  better  surmounted  by  a  siphon. 
A  siphon  consists  of  a  vertically-curved  pipe  with  both  ends  set  in  sumps, 
of  which  the  outlet  sump  must  have  the  lowest  water-level. 

Mine  siphons  are  usually  made  from  welded  iron  pipe  and  water  can 
be  carried  horizontally  in  them  for  considerable  distances  provided  they 
are  tight.  The  limit  of  vertical  lift,  from  surface  of  intake  to  highest 
point  on  the  pipe  of  any  siphon,  is  the  height  of  the  water  barometer 
minus  the  total  loss  of  water  head,  due  to  internal  friction,  etc.,  in  the 
siphon  itself.  This  limit  is  usually  below  26  ft.  Several  rolls  can  be 
passed  by  one  siphon  if  escape  valves  for  air  are  put  on  the  pipe  at  the  high 
point  of  each  vertical  bend.  It  is  also  possible  to  drain  several  sumps  or 
"  swamps  "  along  a  gangway  with  one  siphon,  by  running  a  branch  pipe, 
with  a  valve  on  its  end,  down  from  the  main  siphon  into  each  swamp. 
A  siphon  is  best  rigged  with  a  valve  at  inlet  and  outlet;  and  with  its 
highest  point  joined  by  a  small  pipe,  with  valve,  to  a  water-barrel  from 
which  it  can  easily  be  filled  before  a  run. 

MECHANICAL  UNWATERING 

Apparatus  for  mechanical  drainage  can  be  grouped  into  two  classes. 
First,  those  moving  water  in  buckets,  and  second,  those  moving  water 
through  pipes.  In  the  first  class,  water-cars  are  moved  horizontally  by 
the  same  tractors  as  ore-cars,  while  tanks  or  kibbles  are  hoisted  in  shafts 
or  slopes  by  similar  engines  to  those  used  for  hoisting  ore  in  skips.  The 
second  class  includes  all  types  of  pumps.  The  first  class  is  often  pre- 
ferable for  intermittent  unwatering,  even  if  it  has  a  higher  operating 
cost,  for  where  the  existing  ore-hauling  and  hoisting  equipment  can  be 
utilized  to  move  the  water-buckets,  the  heavy  expense  of  installing 
pumps  is  obviated. 


CHAPTER  VI 
SURFACE  SHOVELING  IN  OPEN  CUTS 

EXAMPLE  1. — MOA  AND  MAYARI  IRON  MINES,  CUBA 

Drag-line  Excavators  on  Shallow  Flat  Placers  Without  Mantle. — No 
soil  surface  exists  over  these  ores;  indeed,  the  ore  itself  is  the  soil,  upon 
which  grow  either  pine  forests  or  a  characteristic  tropical  jungle.  The 
deposites  at  Moa  constitute  a  surface-mantle  varying  in  thickness  from 
a  mere  film  to  121  ft.  and  occupying  an  area  of  60  sq.  miles.  The  area 
of  more  than  8000  hectares  of  ore  drilled  showed  an  average  of  18.83 
ft.;  the  Mayari  deposit  is  a  trifle  thicker  and  shows  an  area  sufficient 
to  contain  more  than  600,000,000  tons  of  commercial  ore.  The  thickness 
of  the  ore-mantle  is  affected  by  local  causes,  assisting  or  delaying  the 
breaking-down  of  the  serpentine  bed-rock  (which  experts  agree  to  be 
the  mother  of  this  ore),  erosion  by  streams,  and  other  causes.  The  ore 
lies  directly  upon  the  serpentine,  and  mining  will  be  somewhat  unfavor- 
ably affected  by  the  fact  that  the  gradation  from  ore  to  rock  is  not  at 
all  regular,  but  very  rough  so  that  in  cleaning  the  bottom  of  an  ore-body 
with  any  sort  of  automatic  machine,  chunks  of  serpentine  are  liable  to 
be  broken  off  and  lifted  with  the  ore,  unless  care  is  constantly  exercised. 
This  ore  is  a  clayey  limonite  containing  nodules  of  magnetite  and  hema- 
tite near  the  surface.  It  shows,  as  chief  constituents, 

Iron  up  to  46  per  cent. 

Alumina  up  to  15  per  cent. 

Silica  up  to  7  per  cent. 

Free  water  up  to  30  per  cent. 

Combined  water  up  to  15  per  cent. 

Prospecting. — By  reason  of  the  character  and  condition  of  these 
ores  exploration  can  be  carried  on  by  a  process  that  is  simple,  accurate, 
rapid,  and  cheap.  Ordinary  2-in.  auger-bits  are  forged  on  one  end  of  4-ft. 
sectional  rods,  and  other  being  fitted  to  receive  a  sleeve-nut,  5  or  6  in. 
long,  into  which  another  4-ft.  section  may  be  screwed.  As  a  hole  is 
driven  down  by  the  auger-bit,  additional  threaded  sections  are  screwed 
on  the  rod,  making  it  any  desired  length.  On  each  end  of  each  rod, 
except  where  the  bit  is  shaped,  to  a  backing-nut  screwed  down  hard, 
in  order  to  prevent  the  rods  from  working  too  tightly  into  the  sleeve-nut 
when  turned  into  the  resisting  ground,  which  would  render  it  difficult 
to  release  quickly.  In  most  cases  ore  can  be  bored  by  this  simple  tool 
with  comparative  ease,  and  when  hard  blocks  and  boulders  are  encoun- 

62 


SURFACE  SHOVELING  IN  OPEN  CUTS 


63 


tered,  they  are  sometimes  cut  by  the  substitution  of  a  cutting  chisel- 
bit  for  the  auger-point :  in  other  cases  the  men  will  move  a  few  feet  away 
and  drive  another  hole,  experience  having  shown  that  a  very  short  distance 
will  usually  be  sufficient  to  avoid  a  boulder.  The  hole  is  started  through 
the  drier  top  soft  ore  or  nodules  on  the  surface,  a  little  water  is  poured  in, 
the  bit  lifted  and  driven  down  by  the  combined  strength  of  two  men,  and 
then  turned  in  the  ore.  The  work  is  a  combination  of  churning  and  bor- 
ing. Every  few  feet  the  tool  is  lifted,  the  ore  adhering  to  the  bit  is 
cleansed  off  by  pressing  a  stick  into  the  point  of  the  bit  and  then  revolving 


FIG.  25. — Drag-line  excavator  at  Mayari,  Cuba. 


the  tool,  and  saved  for  analysis,  and  all  sludge  that  has  collected  above 
the  bit  is  scraped  off.  Were  it  not  for  the  peculiarity  of  this  clayey  ore 
of  standing  without  caving,  this  system  of  drilling  would  be  impossible, 
and  it  would  be  difficult  for  the  engineer  to  follow  and  check  the  depth 
of  holes  by  dropping  down  a  measuring-rod,  or  by  inserting  a  bit  with 
which  to  test  the  material  at  the  bottom.  It  is  not  uncommon  to  check 
grades,  of  properties  previously  drilled,  by  inserting  bits  in  the  old  holes 
and  reaming  out  a  sample  from  the  sides  of  the  hole. 

The  price  paid  the  borers  begins  with  from  1.5  to  2  cents  per  foot 
for  the  first  10  ft.  of  depth,  and  increases  by  the  addition  of  a  like 
sum  per  foot  for  each  succeeding  10  ft.  of  progress  following.  In 
ordinary  ground,  each  borer  will  earn  from  $2.50  to  $3.00  per  day;  in 
other  words,  a  pair  of  borers  will  complete  from  10  to  13  holes,  averag- 


64  MINING    WITHOUT    TIMBER 

ing  20  ft.  deep,  per  day.     One  81-ft.  hole  was  drilled  in  two  long  days 
by  two  men. 

In  no  other  way  it  is  possible  to  explore  such  an  area  except  at  great 
expense  and  in  a  long  time.  No  system  of  tunnels,  pits,  or  other  openings 
is  so  well  suited  to  this  work.  It  is  well  enough  to  sink  pits  occasionally, ' 
to  check  by  actual  observation  certain  facts  that  seem  patent  from  the 
drilling,  or  to  answer  any  questions  that  may  arise. 

To  those  accustomed  to  vein-mines  or  to  the  great  replacement- 
deposits  of  the  Mesabi  iron  range,  borings,  varying  from  100  to  300  m. 
apart  may  seem  utterly  inadequate  to  prove  grades  and  tonnages. 
In  early  examinations  of  the  Mayari  field  original  borings  were  spaced 
every  100  ft.,  but  as  the  work  proceeded  the  ore  was  found  to  be  so 
regular  in  analysis,  texture,  and  thickness  that  holes  were  gradually 
spaced  at  intervals  up  to  and  even  exceeding  1000  ft. 

Excavating. — With  no  over-burden  to  be  removed,  the  deposit  situated 
close  to  the  sea,  with  stream-valleys  cutting  through  the  ore-beds  and 
running  directly  to  deep  water,  and  with  an  average  thickness  suitable 
for  about  one  shovel-cut,  these  ores  may  be  mined  at  low  cost  by  or- 
dinary steam-shovel.  The  drag-line  excavator  is  being  tried  at  Mayari 
(see  Fig.  25)  and  has  advantages  there,  as  the  deposit  of  ore  is  compara- 
tively thin  and  the  floor  quite  rough.  Also,  its  radius  of  action  is  far 
greater  than  that  of  a  steam-shovel,  which  must  be  moved  frequently. 
The  is  no  question  of  the  relative  efficiency  of  the  two  machines  if 
the  shovel  can  get  one  or  two  full  cuts  in  clean  ore,  but  such  opportunities 
are  comparatively  rare. 

* 

EXAMPLE  2 — MESABI  IRON  RANGE,   MINN 
(See  also  Examples  7  and  46.) 

Steam  Shovels  on  Rolling  Lenses  with  Mantle  of  Glacial  Drift.— 
Hidden  as  the  great  deposits  of  the  Mesabi  are  by  a  thick  mantle  of  drift- 
it  is  no  wonder  that  new  bodies  are  even  to-day  being  discovered  after 
16  years  of  active  exploration,  when  it  is  remembered  that  the  produc, 
tive  range  is  100  miles  long  by  one-fourth  to  two  miles  wide. 

The  long,  flat  basins  which  hold  the  ore  are  the  outcome  of  gentle 
folds,  transverse  to  the  range  and  cut  up  into  basins  by  cross-anticlines. 

The  ores  are  mostly  soft,  hydrated  hematites  with  subordinate,  soft 
limonite.  In  the  great,  open  pits  the  occurrence  of  the  ore  in  contin- 
uous beds  of  different  colors  and  grades  is  noticeable. 

The  beds  differ  much  in  texture;  quite  common  are  layers  of  broken 
joint-blocks  of  hard  hematite  from  1/4  in.  to  2  in.  thick,  which 
generally  occur  alternated  with  continuous  layers  of  hematite  sand 
or  dust.  The  pore  space  is  considerable,  which  is  shown  not  only  by 
the  speed  with  which  surface  water  sinks  to  the  drainage  shaft  in  an 


SUEFACE  SHOVELING  IN  OPEN  CUTS  65 

open  pit,  but  also  from  12  cu.  ft.  to  14  cu.  ft.  per  ton  allowed  for  this  ore 
in  place,  as  compared  with  8  cu.  ft.  to  9  cu.  ft.  for  specular  hematite. 

Of  the  coarser  joint-blocks  it  is  possible  to  use  60  per  cent,  to  75  per 
cent,  in  the  blast  furnace  mixture,  while  of  the  dust  33  per  cent,  has  been 
the  practical  maximum.  By  selection  and  judicious  mixing  about  75 
per  cent,  of  the  high-grade  ore  mined  can  be  loaded  as  Bessemer,  and 
much  of  the  remaining  high  grade  is  utilized  for  Bessemer  pig,  after 
mingling  with  lean,  low  phosphorous  ores  from  other  ranges.*  Much 
of  the  lean  non-Bessemer  ore  is  necessarily  removed  in  open-cut  work, 
and  this  has  been  stored  in  stock  piles  for  some  future  use. 

The  Mesabi  ore  deposits  are  enormous,  and  single  bodies  are  known 
to  contain  from  20,000,000  to  40,000,000  tons.  The  deposits  cover 
great  areas,  and,  owing  to  the  drift  mantle  of  10  ft.  to  150  ft.,  many 


-\  A  A 


FIG.  26. — Section  of  ore  body,  Mesabi  range. 

different  adjoining  mines  might  be  on  one  continuous  ore  body  without 
its  being  known.  As  a  rule  the  important  ore  bodies  are  several  acres 
in  area  and  have  a  thickness  increasing  from  the  periphery  to  a  maximum 
of  200  ft.  at  the  center. 

"The  bottom  of  an  ore  body,  where  resting  directly  on  Pokegama 
quartzite,  may  be  smooth,  but  when  on  taconite  it  is  generally  irregular 
and  often  stepped  up  on  the  trough  side.  (See  Fig.  26.)  At  the  bottom 
of  some  large  deposits  are  beds  of  "paint  rock"  and  limonite,  forming  an 
impervious  basement.  In  other  cases  the  good  ore  rests  directly  on 
the  porous  taconite  and  the  basement  of  the  water  circulation  must  be 
looked  for  farther  down,  as,  for  instance,  some  dense  layer  of  quartzite. 

Of  three  prevalent  methods  of  ore  extraction  on  the  Mesabi,  i.e., 
open-cut,  underground  mining  and  surface  milling,  steam  shoveling  is 
easily  first,  underground  mining  second  and  milling  last.  In  1902  the 
second  and  last  systems  produced  46  per  cent,  and  7  per  cent,  of  the 
total  output,  respectively,  but  they  have  since  dwindled  in  relative 
importance,  for  in  1909  they  only  produced  15  per  cent,  of  the  yield  of 
29  1/4  million  tons. 

5 


66 


MINING    WITHOUT   TIMBER 


In  a  new  mine  the  surface  topography  and  the  prospect  drilling 
(which  has  previously  prospected  the  deposit  with  holes  at  200-ft. 
intervals)  will  enable  an  intelligent  choice  of  systems  to  be  made.  With 
suitable  conditions  the  open-cut  method  is  cheapest,  the  milling  second 
and  the  mining  dearest.  The  cost  of  mining  is  about  $1  per  ton,  or' 
twice  as  much  as  average  open-cutting. 

For  open-cutting  a  deep  mantle  may  be  stripped  if  the  ore  beneath 
is  proportionately  thick,  and  the  common  rule  is,  roughly,  that  a  foot 
of  drift  can  be  removed  for  each  foot  of  ore.  With  a  desirable  maximum 
track  grade  of  3  per  cent.,  and  a  possible  one  of  5  per  cent.,  the  proper 


FIG.  27. — Fayal  open  pit. 

layout  of  trackage  to  secure  all  the  ore  is  the  first  consideration.  A 
side-hill  body  means  an  easy  approach;  if,  also,  it  has  an  area  in  proper 
proportions  as  to  total  depth  and  width  so  as  to  allow  for  suitable  benches 
we  have  an  ideal  condition. 

There  must  be  an  available  and  adequate  dump  ground,  and  the 
annual  ore  output  must  be  sufficient  to  cover  the  extra  interest  on  the 
increased  investment  of  this  over  other  systems.  Lastly,  the  lean  layers 
in  the  ore  must  be  harmlessly  situated.  If  one  is  on  the  bottom  it  can  be 
left,  or  if  on  the  top  it  can  be  removed  with  the  strippings;  but  if  inter- 
mediate, so  that  it  cannot  be  separated  in  digging,  it  makes  the  system 
Jess  practical. 


SUKFACE  SHOVELING  IN  OPEN  CUTS 


67 


Drainage. — In  beginning  an  open-cut  mine  a  drainage  shaft  is 
generally  sunk  to  the  lowest  point  of  the  ore  having  two  compartments 
for  a  cage-way  and  a  pump-way.  At  the  bottom  is  placed  a  station  pump 
of  size  proportional  to  the  water  flow,  and  on  the  surface  a  boiler  plant 
of  sufficient  capacity  to  furnish  steam  to  run  a  small  hoist,  a  dynamo 
engine  for  electric  lights  and  the  surface  pumps  which  are  ocasionally 
needed  to  drain  ponds  formed  on  some  impervious  layer  in  the  pit. 

Stripping. — The  general  layout  of  the  digging  operations  is 
planned  entirely  from  the  results  of  prospect  drilling.  There  are  two 
possible  trackage  schemes,  one  a  cut  longitudinally  along  the  long  axis 
of  the  deposit  to  be  worked  outward  to  both  sides,  and  the  other  a  cut 
in  an  elliptical  ring,  from  which  work  proceeds  both  inward  and  outward. 


Scale 
FIG.  28. — Biwabik  open  pit. 

The  topography  determines  the  choice  of  methods,  as  in  early  plans  of  the 
Fayal  and  the  Biwabik  open-cuts  which  incorporate  ideas  from  both  sys- 
tems. (See  Figs.  27  and  28).  Most  stripping  has  been  done  by  con- 
tract, but  recently  some  of  the  operating  companies  have  started  doing  it 
direct  in  order  to  better  utilize  ore-digging  equipment.  The  only  different 
apparatus  needed  for  stripping  is  the  wooden  dirt  car,  which  has  a  shal- 
low body  set  on  a  central  longitudinal  hinge  on  its  truck  so  it  can  be 
dumped  to  either  side.  It  holds  6  cu.  yds.,  has  automatic  couplers  but 
no  brakes,  and  is  made  by  the  Russel  Foundry  Co.  of  Detroit.  Though 
contractors  often  used  a  3-ft.  track  gauge,  it  has  been  fo'und  feasible  to 
use  the  standard  gauge  for  the  dirt  cars  so  as  to  be  uniform  with  that  of 
the  ore  cars  and  shovels.  The  contract  price  for  the  great  Eveleth  bodies 


DO  MINING    WITHOUT    TIMBER 

has  been  30  cts.  a  cubic  yard,  which  for  a  depth  of  10  yds.  would  mean 
nearly  $50,000  an  acre. 

Locomotives. — Steam  locomotives  are  used  which  are  commonly  of  a 
60-ton  size,  with  six  drivers  and  19-in.  by  36-in.  clyinders,  made  by  the 
American  Locomotive  Works.  They  handle  11  of  the  6-yd.  dirt  cars, 
four  25-ton  or  two  50-ton  steel,  bottom-dump  ore  cars.  Some  larger 
locomotives  are  in  use,  and  it  is  planned  at  the  Fayal  mine  to  restrict  the 
60-ton  size  to  service  around  the  shovels.  Longer  trains  will  then  be 
made  up  for  haulage  to  the  terminal  yards  by  100-ton  locomotives. 

Shovels. — For  digging,  the  favorite  shovel  is  the  90-ton  size,  with  a 
dipper  handling  21/2  yds.  of  dirt,  or  41/2  tons  of  ore,  either  of  the 


FIG.  29. — Openpit  work,  Mesabi  range. 

Marion,  Bucyrus  or  Vulcan  make.  This  is  mounted  on  two  4-wheeled 
trucks  on  a  standard-gauge  track  set  20  ft.  from  the  car  track,  which 
allows  a  cut  of  30  ft.  wide  at  the  bottom  (see  Fig.  29).  The  shovel  works 
best  under  a  bench  20  ft.  high,  as  higher  ones  are  apt  to  cause  trouble  by 
caving.  This  makes  it  necessary  to  keep  the  shovel  at  considerable  dis- 
tance from  the  face,  thereby  sacrificing  the  digging  efficiency.  With  work 
well  arranged  this  shovel  can  dig  and  load  150  6-yd.  dirt  cars  or  50  50-ton 
ore  cars  in  10  hours.  Records  much  higher  than  this  can  be  made,  as 
one  shovel  timed  by  the  author  loaded  50  tons  of  ore  in  4  2/3  minutes. 
After  removing  the  bulk  of  the  stripping  the  surface  of  the  ore  body 
must  be  cleaned  before  ore  can  be  dug.  Formerly  this  was  done  by 
hand  shovels  and  barrows,  but  most  of  this  hand  cleaning  has  now  been 


SURFACE  SHOVELING  IN  OPEN  CUTS  69 

superseded  by  the  following  method:  When  the  last  stripping  cut  is 
being  made  a  scoop  scraper  is  chained  to  the  shovel  dipper,  which  drags 
the  dirt  on  the  ore  surface  over  against  the  next  stripping  bench.  In 
1909  nearly  19  million  cu.  yds.  was  stripped  on  the  Mesabi,  of  which 
under  30  per  cent,  was  handled  by  contractors. 

Loosening. — Though  much  of  both  overburden  and  ore  is  soft 
enough  to  be  dug  by  the  shovel  direct,  it  makes  quicker  work  to  have  the 
benches  first  loosened  by  explosives.  The  holes  are  bored  in  gravel  or  ore 
with  hand  churn  drills  to  the  same  depth  as  the  height  of  the  shovel  bench. 
They  are  placed  15  ft.  to  20  ft.  apart  along  a  bench  and  staggered,  two 
abreast.  The  drills  are  of  1-in.  to  3  1/4-in.  octagonal  steel,  with  1  1/2-in. 
chisel  points,  and  are  operated  by  two  or  four  men,  according  to  depth 
reached.  In  ore,  24  ft.  per  man  is  drilled  in  10  hours,  but  in  the  drift, 
with  its  boulders,  only  10  ft. 

The  drills  are  lifted  by  a  movable,  steel  cross-piece,  one  for  each  pair 
of  men,  held  to  the  shank  by  wedges.  A  small  intercepting  boulder  may 
be  dislodged  by  a  squib,  but  the  larger  ones  are  drilled.  The  finished 
hole  is  squibbed  several  times  with  60  per  cent,  dynamite  to  make  a 
chamber  for  three  to  eight  kegs  (25  Ibs.)  of  1/4-in.  black  powder,  which 
is  not  loaded  until  the  hole  has  been  dried  by  pouring  in  sifted  sand. 

Spoil-banks. — A  part  of  many  of  the  open-cut  mines  was  originally, 
or  is  still,  worked  by  caving,  and  the  resultant  surface  depressions  can 
often  be  utilized  as  convenient  dump  holes  for  stripping.  The  large 
area  of  wild,  rolling  land  around  the  ore  bodies  makes  it  easy  to  find  a 
dump  for  the  balance  without  too  long  a  haul.  The  common  method  is 
to  lay  the  switch  on  a  side  hill  and  dump  the  dirt  down  hill  until  the  dump 
becomes  high  enough  to  move  the  track  sideways  to  the  edge,  where  the 
dumping  process  is  repeated.  To  keep  the  dirt  from  clogging  the  track 
it  is  customary  to  level  the  dump  by  a  plow-like  scraper  with  a  V-shaped 
nose,  mounted  on  a  heavy  truck  and  adjusted  for  different  heights. 

At  Coleraine  the  Oliver  Co.  has  adopted  a  new  scheme  for  dis- 
posing of  dirt  from  the  Canisteo  mine.  The  initial  dump  was  raised 
by  degrees  to  a  height  of  50  ft.  along  the  shore  of  Trout  lake,  and  a 
trestle  resting  on  piles  was  then  built  to  support  the  track  on  the  edge  of 
the  dump  facing  the  lake.  From  this  trestle  the  dirt  cars  are  dumped 
sideways  into  the  lake,  and  the  dirt  is  kept  from  accumulating  by  the  use 
of  water  jets  played  upon  it  from  above.  The  scheme  saves  removal  of 
the  terminal  track,  the  only  extra  expense  being  for  wash  water,  which 
is  raised  by  a  special  pump  located  below  the  dump. 

Labor. — In  open-cutting  the  following  force  is  employed:  For  shovel 
9  (runner,  craneman,  fireman,  four  pitmen  and  two  track  cleaners); 
for  trains,  3  (engineer,  fireman  and  brakeman) ;  then  there  are  the  blasting 
and  dump  bosses  and  their  helpers,  also  the  engineer  and  pumpman  at 
the  drainage  shaft.  The  general  force  comprises  the  superintendent, 


70  MINING    WITHOUT   TIMBER 

day  foreman  and  night  foreman,  the  sampler,  surveyors,  assayers, 
accountants  and  clerks. 

Shops. — For  keeping  the  heavy  machinery  up  to  its  work  extensive 
repair  shops  are  maintained  by  the  different  operators.  For  instance, 
the  Oliver  Company 's  mines  at  Hibbing  have  a  blacksmith  shop  with  six 
forges  and  a  steam  hammer,  a  machine  shop  with  planer,  boring  mill, 
drill  and  wheel  presses,  lathe,  shaper,  etc.,  large  enough  to  handle 
any  heavy  repairs  and  renewals  for  shovels  or  locomotives.  There  is 
also  a  foundry  for  medium-sized  castings;  the  large  castings  are  shipped 
in  the  rough  and  finished  to  dimensions  at  the  mine  machine  shop. 

Prospecting. — The  early  mines  of  the  Lake  Superior  iron  ranges 
were  started  on  ore  out-crops  showing  through  the  glacial  drift,  which 
had  been  followed  up  by  test-pitting.  In  the  last  few  years,  churn 
drilling  has  been  in  vogue  in  drift  and  soft  ore,  while  in  hard  ore  or  rock, 
the  diamond  drill  is  used. 

On  the  Mesabi  range,  prospecting  has  been  very  systematic,  and  it 
is  estimated  that  over  30,000  holes  have  been  drilled  from  the  surface. 
The  unit  of  area  is  the  40-acre  lot,  and  (where  not  adjoining  ore-bearing 
ground)  if  nothing  is  found  by  drilling  one  hole  near  the  center,  and  one 
at  each  of  the  corners,  the  lot  is  considered  as  barren.  The  surface 
boundaries  of  the  iron  formation  have  now  been  so  well  mapped  by  the 
government  geologist  that  there  is  no  further  excuse  for  wild-catting 
on  impossible  areas.  If  ore  is  penetrated,  a  hole  is  put  down  at  each 
200-ft.  interval,  and  records  are  kept  to  determine  not  only  the  outlines 
of  the  ore  body,  but  also  the  depth  and  assay  of  each  of  the  rock  and  ore 
strata  penetrated. 

The  Mesabi  prospecting  is  started  with  a  churn  drill  which  is  of  the 
portable  type  run  by  a  gasoline  or  steam  engine.  A  boring  sample  is 
taken  for  every  5  ft.  of  depth  and  assayed  for  SiO2,  Fe,  Mn,  P,  and  Al. 
(The  Al  assay  has  lately  been  added  to  determine  if  the  sample  is  from 
aluminous  paint  rock  or  from  the  ore  proper.)  The  bit  is  only  wide 
enough  to  go  inside  a  3-in.  welded  pipe  casing,  which  is  kept  within 
5  ft.  of  the  bottom  of  the  hole.  When  5  ft.  below  the  bottom  of  casing 
has  been  drilled,  and  the  sample  taken,  the  casing  is  forced  down  to  the 
bottom.  The  sample  it  obtained  by  running  the  sludge  through  two 
settling  boxes,  from  which  the  slime  is  saved  and  sent  to  the  assay  labor- 
atory, where  it  is  dried  and  cut  down  and,  after  taking  enough  for  analy- 
sis, about  1  Ib.  is  kept  in  a  tubular  box  for  reference. 

On  reaching  the  hard  schist,  called  Taconite  (see  Fig.  26),  below  an 
ore  body,  the  hole's  casing  is  seated  there  and  the  churn  replaced  by  a 
1  1/4-in.  diamond  drill,  taking  a  7/8-in.  core.  If  ore  is  again  penetrated, 
the  hole  is  enlarged  sufficiently  to  let  the  casing  descend.  This  is  done 
by  lifting  the  casing  off  the  rock  and  then  blasting  the  top  of  the 
1  1/4-in.  hole  with  a  stick  or  two  of  60  per  cent,  dynamite,  fired  elec- 


SUKFACE  SHOVELING  IN  OPEN  CUTS  71 

trie  ally.  The  pieces  are  then  cleaned  out  by  the  churn-bit  and  sand 
pump,  and  the  blasting  repeated  until  the  hole  is  enlarged  sufficiently 
to  drop  the  casing  down  to  the  ore.  The  churn  drill  is  then  used  until 
the  bottom  of  the  ore  is  reached.  This  alternating  churn  and  diamond 
drilling  is  kept  up  until  the  basal  quartzite  is  reached,  beneath  which  no 
ore  is  ever  found. 

The  whole  diamond  drill  core  is  not  saved,  but  only  1  ft.  or  so  of  rep- 
resentative rock  for  each  5  ft.  of  core,  and  this  is  stored  in  core  boxes, 
packed  in  special  cases  along  with  the  churn  drill  samples,  in  a  fire-proof 
vault  in  the  office  building.  The  drilling  is  done  for  the  mining  com- 
panies by  contractors  at  a  cost  for  churn  drilling  of  50  cents  to  $1  a  foot. 
The  diamond  drilling  costs  about  twice  as  much  per  foot,  but  it  is  eco- 
nomical in  taconite,  owing  to  the  slowness  of  the  churn.  To  prevent 
the  contractors  using  the  faster  and  more  profitable  (to  themselves) 
diamond  drill  in  soft  ore,  they  must  show  rock  cores  for  all  such  work. 

Sampling  and  Assaying. — One  of  the  most  complete  systems  is  in 
vogue  at  the  Oliver  Co.'s  mines  on  the  Mesabi.  For  underground  work, 
a  grab  sample  is  taken  from  each  skip  hoisted,  and  a  day's  hoist  is 
assembled  and  cut  down.  Both  underground  and  open-pit  faces  are 
sampled  by  the  bosses  (by  grooving)  frequently  enough  to  insure  intelli- 
gent stoping. 

For  open-pit  shipments,  a  separate  sample  it  taken  from  each  10  cars 
of  30  to  50  tons  each.  This  is  done  by  grabbing  10  to  12  handfuls  off 
the  surface  of  each  car,  from  spots  spaced  equally  along  diagonals  or 
two  longitudinal  parallel  lines.  The  assembled  samples  from  10  cars  are 
then  reduced  in  the  pit  by  the  sampler  to  10  Ibs.  (by  coning  and  quarter- 
ing), and  sent  in  a  sack  to  the  assay  laboratory. 

In  the  laboratory  the  10-lb.  sample  it  first  bucked  by  hand  to  pass 
a  2-mesh  sieve,  and  then  cut  down  on  oilcloth  by  coning  to  1  kg.,  and 
the  balance  used  for  determining  moisture.  The  1  kg.  is  crushed  in  a 
laboratory  jaw  crusher  and  quartered  to  100  grams,  which  is  finally 
dried,  bucked  by  hand  to  100-mesh,  and  stored  in  a  sample  bottle. 

All  shipments  are  analyzed  for  Si02,  Fe,  Mn,  and  P,  and  the  drillings 
for  Al  in  addition.  , 

The  sampling  and  analysis  of  a  Mesabi  shipment  must  be  completed 
before  the  train  reaches  Two  Harbors,  which  often  allows  only  four  hours. 
The  result  can  then  be  telephoned  to  the  dock  master  and  he  can  assign 
the  shipment  to  its  appropriate  pocket  on  the  dock.  In  this  way,  cars 
of  different  grades  to  make  a  desired  compound  can  be  dumped  into  the 
same  pocket,  and  by  the  time  the  furnace  is  reached,  the  several  tran- 
shipments will  have  produced  a  uniform  mixture. 

Surveying. — For  the  Mesabi  open  pits,  the  Oliver  Co.  pursues  the 
following  survey  system:  Before  the  contractors  begin  stripping,  the 
whole  area  is  blacked  out  in  100-ft.  squares  and  stakes  set  across  the 


72  MINING    WITHOUT    TIMBER 

deposit  on  the  100-ft.  lines  at  20-ft.  intervals.  The  level  of  each  station 
is  then  taken,  and,  by  holding  a  tape  between  neighboring  stakes  to  make 
the  longitudinal  20-ft.  marks,  the  remaining  levels  can  be  observed  at 
the  corners  of  each  20-ft.  square.  When  the  drift  has  been  removed 
down  to  the  ore,  the  new  level  of  each  20-ft.  point  is  observed,  and  then 
the  total  volume  of  stripping  can  be  accurately  calculated.  Recently 
the  company  has  started  doing  some  stripping  on  company  account, 
and  here  levels  are  taken  only  at  40-ft.  intervals. 

Subsequently  the  ore  removed  during  any  given  period  can  be  easily 
calculated  by  securing  new  elevations  of  the  same  points  used  in  the 
stripping  survey.  The  ore  reserves  were  formerly  computed  from  the 
drill-hole  records  by  treating  each  body  as  a  whole.  Owing  to  the  fact 
that  the  Mesabi  ore  runs  in  thick  layers  of  large  area,  but  differing  analy- 
ses, it  has  been  found  possible,  with  the  same  drill  records,  to  estimate 
the  volume  of  each  ore  layer  separately,  and  this  has  made  the  results  of 
the  reserve  calculations  much  more  instructive  and  useful. 

EXAMPLE  3. — UTAH  COPPER  MINE,   BINGHAM,  UTAH 
(See  also  Examples  37,  41  and  43.) 

Steam  Shovels  on  Steep  Sidehill  Lenses  with  Rock  Capping. — The 
Bingham  orebody  seems  to  owe  its  existence  to  the  impregnation  of  the 
shattered  zones  in  a  monzonite-porphyry  intrusion  that  was  forced  up 
vertically  through  the  surrounding  quartzites.  The  mineralization  is 
not  confined  entirely  to  the  porphyry,  for  in  the  Ohio  Company's  ground 
and  in  the  Starless  group  shattered  areas  in  the  quartzite  itself  have 
been  strongly  mineralized.  In  the  course  of  time  this  porphyry  ore  suc- 
cumbed to  erosion  and  oxidation  more  rapidly  than  the  less  shattered 
quartzites,  and  so  it  now  forms  the  bottom  of  a  gulch  and  extends  up  to 
the  top  of  the  divide  between  Main  Bingham  and  Carr  Fork  canyons. 

In  the  upper  part  of  the  sulphide  ore  at  Bingham  the  copper  in  the 
ore  has  been  leached  from  the  walls  of  the  seams,  while  between  the 
seams  in  the  unreached  center  of  the  porphyry  the  ore  is  unaffected 
and  assays  fairly  well  in  copper.  This  fact  would  indicate  that  the 
ore  approached  the  present  richness  before  secondary  enrichment.  Some- 
what deeper  the  ore  becomes  richer  and  then  in  turn  drops  in  grade 
until  it  falls  below  workable  value.  The  principal  copper  mineral  in  the 
sulphide  ore  is  chalcocite,  disseminated  in  extremely  fine  particles. 

Leaching  has  progressed  concomitant  with  erosion,  and  so  the  ore- 
body  is  covered  with  a  capping  of  oxidized  porphyry  from  which  some 
of  the  copper  has  been  carried  away  in  solution.  Still  much  of  this 
carries  well  above  1  per  cent,  and  in  places  as  much  as  2  and  even  2.5 
per  cent.  This  copper  of  course  occurs  in  oxidized  form,  and  therefore 
is  not  so  amenable  to  concentration  as  the  sulphide  ore.  At  present 


SURFACE  SHOVELING  IN  OPEN  CUTS  73 

all  the  capping  is  looked  upon  as  waste  no  matter  how  much  copper  it 
may  contain.  This  line  of  demarkation  between  capping  and  ore  is  not 
distinct,  but  it  roughly  parallels  the  surface,  causing  difLculty  in  steam- 
shoveling  the  ore  on  account  of  its  slope,  which  is  too  flat  to  permit 
the  loosened  capping  to  run  down  to  the  shoveling  terraces  and  too  steep 
to  allow  easy  access  without  many  terraces.  Ihe  capping  averages  70  ft. 
thick. 

Owing  to  the  cost  of  underground  mining,  the  Utah  Copper  company 
now  obtains  most  of  its  ore  by  means  of  steam  shovels.  But,  owing  to 
the  fact  that  the  property  is  cut  in  two  by  a  gulch  whose  sides  approxi- 
mate a  slope  of  30  deg.,  the  orebody  is  not  especially  adapted  to  steam- 
shovel  mining.  Indeed,  because  of  the  rough  topography,  already  16 
miles  of  track  have  been  necessary — even  when  a  switch-back  system 
for  entering  the  terraces  and  grades  varying  from  2  per  cent,  to  6  per 
cent,  are  used.  Besides,  on  account  of  the  slope  of  the  gulch  and  the 
line  of  merger  between  capping  and  ore,  there  is  much  mixing  of  over- 
burden and  ore.  In  addition,  the  disposal  of  this  capping  has  necessitated 
the  purchase  of  dump  ground  involving  a  maximum  haul  of  1  1/2  miles. 

The  difference  in  elevation  from  the  creek  level  to  the  highest  ore  is 
900  ft.,  so  that  nine  terraces  have  been  necessary  and  a  tenth  is  now  being 
surveyed.  The  elevation  of  these  are:  /,  6825  ft.;  H,  6750;  G,  6687;  F, 
6600;  E,  6536;  C,  6415;  and  A,  6340  ft.  Stripping  tracks  are  placed  at 
whatever  elevation  the  lay  of  the  ground  dictates.  Stripping  is  done  on 
all  of  these  levels,  but  all  the  ore  above  the  F  line  is  to  be  shoveled 
down  to  the  A  pit  and  loaded  there,  it  being  cheaper  to  handle  the  ore 
twice  than  to  load  it  into  cars  on  the  upper  terraces.  Indeed,  in  A  pit 
much  of  the  ore  is  being  shoveled  from  a  bank  230  ft.  high  (see  Fig.  30), 
a  condition  quite  dangerous  for  the  shovelmen. 

Both  the  ore  and  the  capping  require  blasting  in  order  to  loosen  them 
for  the  steam  shovel.  This  is  done  in  three  different  ways,  according 
to  the  varying  conditions.  In  case  there  is  sufficient  room  a  "gopher" 
is  used,  consisting  of  a  drift,  2  1/2  ft.  square,  driven  into  the  bank  for 
30  to  45  ft.  and  with  a  cross  drift  about  15  ft.  long,  driven  to  each  side 
at  its  end.  This  is  loaded  with  black  powder,  some  dynamite  being  used 
to  explode  the  charge.  In  case  that  there  is  not  quite  so  much  room 
for  blasting,  churn-drill  holes  are  resorted  to.  In  drilling  these  holes 
No.  3  Keystone  churn  drills  are  used.  There  are  two  of  these,  one 
working  on  capping  and  the  other  on  ore.  The  holes  have  a  diameter 
of  6  in.  and  are  cased  generally  only  down  to  solid  rock.  A  crew  consists 
of  a  driller  at  $5  per  12-hr,  shift,  a  tool  sharpener,  as  the  helper  is  called, 
at  $4,  and  a  man  to  get  coal,  water,  etc.,  at  $1 . 75,  and  drilling  and  loosen- 
ing costs  about  3/4  cents  per  ton. 

The  depth  of  these  holes  varies  considerably,  as  well  as  their  spacing 
from  the  bank,  but  when  a  hole  is  to  be  drilled  50  ft.  deep  it  is  generally 


74 


MINING    WITHOUT    TIMBER 


placed  so  as  to  have  a  burden  of  35  ft.  at  its  bottom.  In  some  cases 
holes  85  ft.  deep  have  been  used;  these  are  loaded  generally  in  three, 
places,  the  bottom  only  being  sprung.  In  case  the  bank  is  low,  hori- 
zontal holes,  drilled  with  a  3  1/2  in.  piston  machine  to  a  depth  of  25  to 
26  ft.,  are  used.  These  are  sprung  2  or  3  times  so  that  they  will  hold' 
three  boxes  of  30  per  cent,  dynamite.  Almost  all  blasting  is  done  with 
a  battery,  and  three  or  four  caps  are  used.  Still  misfires  occur  even  with 
these  precautions,  but  they  are  rare.  Most  of  the  capping  and  ore  breaks 
quite  fine,  owing  to  the  fracturing  that  has  occurred  throughout  the  mass 


FIG.  30. — Shoveling  230-ft.  bank,  Utah  Copper  mine. 

of  porphyry,  but  occasional  bulldozing  of  boulders  is  necessary.  This  is 
cheaper  than  block-holing  the  boulders. 

After  each  blast  the  bankmen,  by  means  of  ropes,  climb  down  the 
sides  and  dress  down  the  bank.  These  men  also  do  all  blasting  and 
gophering.  Their  work  is  especially  dangerous,  and  they  are  paid  $3 . 75 
for  10-hr,  shifts.  The  difficulty  of  their  work  can  be  seen  in  Fig.  31. 
The  men  work  with  a  rope  near  at  hand  to  grab  in  emergencies. 

The  banks  in  capping  are  carried  at  an  angle  of  about  40  deg.,  as 
this  is  the  best  slope  for  working  it;  the  ore  bank  230  ft.  high  is  carried 
at  a  somewhat  steeper  angle.  The  capping  is  shoveled  into  4  1/2-cu.  yd. 


SURFACE  SHOVELING  IN  OPEN  CUTS 


75 


dump  cars  and  run  in  trains  to  the  waste  dumps,  while  the  ore  is  shoveled 
directly  into  50-ton  railroad  cars  ready  for  hauling  to  the  mill  at  Garfield. 
The  routine  of  handling  the  shovels  is  quite  similar  to  that  in  Example 
2,  and  so  will  not  be  described.  A  65-lb.  rail  is  used  and  a  standard 
broad-gauge  track;  in  advancing  the  shovel  6-ft.  lengths  of  track  joined 
by  fish  plates  are  used.  A  shovel  crew  consists  of  a  shovelman  at  $175 
per  month,  a  craneman  at  $125  per  month,  a  fireman  at  $2.50  a  day, 
and  six  pitmen  at  $1 . 75.  Four  of  the  pitmen  work  on  the  track  and 
odd  jobs,  while  two  tend  the  jack  screws  and  help  advance  the  shovel. 


FIG.  31. — Trimming  bank,  Utah  Copper  mine. 

The  shovels  work  from  60  to  65  per  cent,  of  the  time.  The  rest  of 
the  time  is  taken  up  mainly  in  waiting  for  cars,  blasting  and  other 
similar  delays.  As  yet,  few  shovels  have  been  buried  by  caves,  and  there 
are  not  many  break-downs.  The  company  has  10  shovels;  eight  weigh 
95  tons  and  have  3  1/2-cu.  yd.  dippers,  while  two  weigh  65  tons  and 
have  3-cu.  yd.  dippers.  Seven  are  Marion  shovels,  two  Bucyrus  and  one 
Vulcan,  the  latter  being  one  of  the  first  bought.  The  shovels  work 
two  10-hr,  shifts.  The  large  shovels  consume  about  21/2  tons  of  coal 
per  shift  and  the  smaller  ones,  2  tons.  Seven  of  the  shovels  are  working 
on  capping  all  the  time;  two  shovels  on  ore,  and  one  partly  on  ore  and 


76  MINING    WITHOUT    TIMBER 

partly  on  capping.  About  6500  tons  of  ore  (2  1  /4  to  7  per  cent,  moisture) 
and  8000  to  10,000  tons  of  capping  were  mined  in  a  day  in  June,  1909, 
with  about  650  men  at  this  work  which  cost  19  cents  per  ton  of  ore  and 
34  cents  per  cu.  yd.  of  capping  removed.  In  1911  with  19  shovels,  ore 
and  capping  are  handled  at  the  rate  of  50,000  tons  daily. 

EXAMPLE  4.     NEVADA  CONSOLIDATED  MINES,  ELY,  NEVADA 
(See  also  Example  44.) 

Steam  Shovels  on  Rolling  Lenses  With  Rock  Capping. — The  ore 
occurrence  at  Ely  is  not  nearly  so  similar  to  that  at  Bingham  as  man}' 
seem  to  think.  At  Bingham  the  intrusion  of  monzonite  porphyry  was 
laccolithic  in  character,  while  at  Ely  the  intrusion  was  more  in  the  nature 
of  a  dike.  It  cuts  persistently  across  the  bedding  planes  of  the  different 
horizons  of  the  limestone  country. 

The  whole  area  has  been  subject  to  much  mineralizing  action,  and 
the  monzonite  has  been  much  kaolinized  in  places  so  that  it  is  much 
weaker  than  at  Bingham.  In  fact  most  drifts  in  the  Ely  monzonite 
have  to  be  timbered,  while  at  Bingham  the  drifts-  stand  without  any 
need  of  support. 

The  orebodies  at  Ely  are  due  essentially  to  secondary  enrichment, 
as  is  clearly  indicated  by  the  sudden  change,  within  a  vertical  distance 
of  5  ft.  from  capping  carrying  only  a  trace  of  copper  to  ore  carrying 
1.5  per  cent,  copper.  This  secondary  enrichment  occurred  at  water 
level,  and  so  the  top  of  the  orebodies  is  fairly  flat,  rarely  undulating- 
through  a  vertical  height  as  great  as  30  ft.  in  the  orebodies  near 
enough  to  surface  to  be  worked  by  steam  shovels,  but  being  somewhat 
greater  in  the  deposits  covered  by  capping  30  ft.  or  more  thick. 

This  flat  character  of  the  top  of  the  orebody,  the  fact  that  it  bears 
no  relation  to  the  undulations  of  the  surface  immediately  above  it,  and  the 
general  rolling  character  of  the  surface  at  the  places  where  the  capping 
is  thin  enough  to  permit  economic  stripping  with  steam  shovels,  makes 
Ely  an  ideal  camp  for  the  use  of  shovels. 

The  depth  of  the  capping  varies  throughout  the  district,  being  only 
30  ft.  thick  in  some  places  and  over  700  ft.  in  others,  depending  partly 
upon  the  height  of  the  local  ground-water  level  when  that  particular 
orebody  was  formed,  but  mainly  upon  the  amount  of  erosion  that  has 
occurred  subsequently  to  the  formation  of  the  orebody. 

The  principal  ore  mineral  is  chalcocite,  but  some  bornite  and  some- 
melaconite  is  also  found.  The  original  mineral  seems  to  have  been 
chalcopyrite.  Owing  to  the  important  influence  of  secondary  enrich- 
ment, most  of  the  copper  minerals  occur  along  the  fracture  planes  and  in 
the  more  porous  rock. 

The  orebodies  occur  where  there  are  shattered  zones  in  the  porphyry 
and  also  where  kaolination  has  allowed  surface  water  to  filter  easily 


SURFACE  SHOVELING  IN  OPEN  CUTS  77 

through  the  porphyry.  Because  of  the  direct  connection  between 
facility  for  leaching  and  richness  of  underlying  orebody,  it  would  appear 
that  there  is  much  less  likelihood  of  finding  orebodies  under  the  lime- 
stone capping  than  where  the  porphyry  comes  to  the  surface. 

The  Nevada  Consolidated  Copper  company  has  three  No.  5  Keystone 
churn  drills  busy  in  prospecting — one  at  the  Boston-Montana-Liberty 
orebody  and  two  at  the  Ruth,  all  in  ore.  The  holes  are  placed  200  ft. 
apart  and  are  sampled  in  5-ft.  sections.  The  entire  pulp  is  dried 
without  decanting  any  of  the  water,  a  sample  from  a  5-ft.  depth 
generally  requiring  four  tubs.  After  drying,  all  the  pulp  is 
mixed  and  then  quartered  down.  The  depth  of  capping  at  the 
Ruth  is  300  ft.,  and  at  the  Boston  and  Montana,  from  80  to  100  ft. 
Few  of  the  holes  require  casing.  At  the  Ruth  the  holes  vary  from 
500  to  700  ft.  in  depth,  while  on  the  Boston  and  Montana  they 
are  only  200  to  250  ft.  deep.  In  holes  deeper  than  350  ft.  about 
25  ft.  are  drilled  in  12  hours,  but  in  the  shallow  holes  more  than  110  ft. 
have  been  drilled  in  12  hours,  the  amount  ordinarily  varying  between  40 
and  60  ft.  in  holes  200  ft.  deep.  In  deep  holes  the  first  casing  used  is  a 
7  5/8-in.,  next  61/4  in.  and  finally  if  required  4  1/4-in.  casing,  but  in 
shallow  holes  or  in  ground  the  nature  of  which  is  known  generally 
an  8-in.  hole  is  started.  The  cost  of  drilling  in  the  Ely  district  varies 
considerably,  but  at  one  mine  churn  drilling  cost  $1.87  a  foot  for  44 
holes  averaging  250  ft.  in  depth.  This  included  the  cost  of  changing 
set-ups,  pulling  casing,  repairs,  lost  tools,  etc.  Complete  records  of  the 
character  of  the  ground  penetrated,  the  amount  of  depth  drilled  each 
shift,  and  the  cost  are  kept.  The  holes  are  plotted,  and  on  sections  the 
assay  of  each  5-ft.  section  is  recorded.  A  double  record  is  kept  of  each 
hole.  From  these  sections  it  is  possible  to  estimate  the  respective  ad- 
vantages of  stea  m-shovel  and  of  underground  mining. 

The  Copper  Flat  orebody  is  admirably  adapted  to  steam-shovel 
work  as  the  capping  is  only  from  35  to  160  ft.  deep  (the  average  being 
about  90  ft.),  the  orebody  210  ft.  deep,  and  the  topography  gently  rolling. 

For  the  benc  h  a  vertical  height  of  50ft.  has  been  found  most  admirable 
for  safety  and  for  intensive  shoveling,  while  with  a  horizontal  width  of 
50  ft.  the  loadir  g  track  is  not  so  likely  to  be  buried  by  the  blast,  thus 
avoiding  serious  and  costly  delays.  Of  course,  higher  banks  can  be  car- 
ried, a  height  of  200  ft.  or  more  being  possible,  but  this  would  necessitate 
tunnel-blasting,  requiring  a  heavy  tonnage  of  explosives,  a  much  wider 
bench,  and  more  care  to  avoid  exposing  the  shovel  and  crew  to  danger 
from  a  treacherous  bank. 

A  study  of  an  actual  section  (Fig.  32),  through  a  bank  of  the  steam- 
shovel  ore-pit,  shows  that  the  slope  between  the  benches  from  the  upper 
edge  to  the  toe  of  the  talus  below  is  a  little  greater  than  1  to  1,  varying 
between  1 ,04  ami  t  -  43,  or  an  average  of  1 . 18  to  1.  The  talus,  or  broken 


78 


MINING    WITHOUT    TIMBER 


FIG.  32. — Bench-diagram,  easy  slope,  Nevada  Consolidated  open  pit. 


7/S0 


FIG.  33. — Bench  diagram,  steeper  slope,  Nevada  Consolidated  open  pit. 


SURFACE  SHOVELING  IN  OPEN  CUTS 


79 


material,  that  has  become  loose  and  has  fallen  to  the  bench  below,  as 
shown  in  the  diagram,  will  always  be  found  at  the  foot  of  a  bank,  the 
quantity  varying  in  amount  according  to  the  condition  of  the  standing 
ground.  In  the  analysis  of  a  steeper  section,  as  shown  in  Fig.  33,  the 
ratios  of  the  corresponding  banks  are  somewhat  less,  varying  between 
0 . 8  and  1 . 18,  or  an  average  for  the  four  banks  of  0 . 99  to  1.  The  steepest 
bank  in  the  shovel-pit  is  one  in  the  zone  of  sulphide  ore,  shown  above 
in  Fig.  33,  52.9  ft.  high,  and  standing  at  a  ratio  of  0.6  to  1.  This  bank 
is  freshly  cut  and  will  stand  at  this  ratio  for  only  a  short  time,  when 
disintegration  will  cause  it  to  crumble.  It  will  be  seen,  then,  that  an 
average  slope  for  this  height  of  bank  and  material  will  average  closely 
the  ratio  of  1  to  1,  a  little  steeper  in  the  zone  of  sulphides  and  a  little 
flatter  in  the  oxidized  material  above. 

The  general  slope  from  the  bottom  to  the  upper  edge  of  the  excavation 
is,  of  course,  much  larger  since  the  added  width  of  the  bench  on  which 
the  shovel  operates  nearly  doubles  the  horizontal  distance.  In  Fig.  32 
this  ratio  over  all  from  the  edge  of  the  top  to  the  toe  of  the  bottom 
bench,  will  be  seen  to  be  1.92  to  1,  and  in  Fig.  34  1.76  to  1  for  four 
benches.  With  the  addition  of  more  benches  the  ratio  will  not  remain 
the  same,  but  will  grow  larger  by  a  decreasing  amount.  Using  the  ideal 
section  (Fig.  29),  with  1  to  1  slopes,  50-ft.  benches  and  50-ft.  heights, 
the  table  and  formula  below  are  suggested  by  E.  E.  Barker. 


Nunber 

Horizontal 

Vertical 

Ratio  of 

Ratio 

of  benches. 

distance. 

distance. 

slope. 

Difference. 

2 

150 

100 

1.50  to  1 

3 

250 

150 

1  .  66  to  1 

0.16 

4 

350 

200 

1.75  to  1 

0.09 

5 

450 

250 

1.80  to  1 

0.05 

6 

550 

300 

1.83  to  1 

0.03 

From  this  table  the  formula  below  is  deduced: 
na+(n-l)b 


-or.  substituting 
nc 

4X50  +  3X50 


350 


S  = 


Where  S  equals  the  slope  ratio;  a  equals  the  base  of  the  individual 
slope  triangle;  b  equals  the  width  of  bench;  c  equals  the  vertical  height 
of  bank;  n  equals  the  number  of  benches. 

Again,  with  c  =  60  ft.;  a  =  50  ft.;  6  =  60  ft.;  w  =  6,  we  have 
6X50  +  5X60     600 


80 


MINING    WITHOUT    TIMBER 


In  the  selection  of  the  width  of  bench,  the  deciding  factor  is  the  slope 
taken  by  the  blasted  bank.  Ordinary  broken  material  will  repose  at  a 
slope  of  about  1  1/2  to  1,  but  the  impetus  given  the  broken  rock  in  a 
blast  usually  causes  the  slope  to  form  at  approximately  the  ratio  of  2 
to  1.  Since  the  drill-holes  are  placed  about  10  ft.  from  the  edge  and  the 
blast  loosens  the  ground  for  about  10  ft.  more,  the  bank,  when  blasted, 
lying  at  a  slope  of  2  to  1  assumes  the  position  of  the  dotted  line  at  the 
top  of  Fig.  34,  still  leaving  20  ft.  clear  on  the  bench  below,  which  gives 
ample  room  to  safely  accommodate  the  loading  track  without  danger 
of  being  covered.  From  the  data  at  hand  and  with  conditions  as  given 


APPROXIMATE  POS/r/OH 
rOff  BUSTED  BANK 
Slope  2  to/. 


FIG.  34. — Bench-diagram,  ideal  slope,  Nevada  Consolidated  open  pit. 

above,  the  section  with  the  minimum  bench  and  the  maximum  height 
and  slope  for  economical  operation,  is  the  one  shown  in  Fig.  34,  with  a 
general  slope  of  1.75  to  1,  or  a, corresponding  ratio  for  the  number  of 
terraces  required.  Already  the  steam-shovel  operations  cover  an  area  of 
many  acres.  The  pit  is  roughly  oval  in  shape  and  the  tracks  are  extended 
in  ovals  around  the  sides,  so  that  only  in  starting  a  terrace  is  it  neces- 
sary to  load  the  cars  singly. 

LOOSENING  ORE  FOR  SHOVELS 

The  ore  and  the  capping  require  blasting  to  loosen  it,  but  on  the 
east  side  the  ground  is  almost  soft  enough  to  shovel  without  blasting. 
Only  churn  drills  are  used  in  preparing  the  bank  for  blasting.  After 
the  ground  along  the  terrace  has  been  roughly  evened,  so  as  to  permit 


SURFACE  SHOVELING  IN  OPEN  CUTS  81 

the  drills  to  move  readily  from  one  set-up  to  another,  drilling  begins. 
These  holes  are  drilled  about  10  ft.  deeper  than  the  steam-shovel  terrace 
so  that  the  bottom  will  surely  be  loosened.  The  holes  are  placed  so  as 
to  have  a  burden  of  from  25  to  50  ft.  of  ground  at  their  bottom,  according 
to  the  nature  of  the  ground,  and  the  holes  are  spaced  approximately 
the  same  distance  apart  as  they  have  burden  on  them.  At  present  there 
are  seven  distinct  kinds  of  ground  to  be  blasted.  This  variability  also 
effects  the  loading  of  the  hole.  The  hole  is  first  sprung  or  chambered, 
3  or  4  times  with  40  to  100  Ib.  of  40  per  cent,  dynamite,  then  the  hole 
is  loaded  with  from  750  to  2000  Ib.  of  explosive.  Several  different 
grades  of  explosives  are  used — in  soft  ground  Dupont  FF  black 
powder,  in  harder  ground  Champion  powder  (a  mixture  between 
black  powder  and  dynamite),  and  is  still  harder  ground  40  per  cent, 
dynamite  (in  winter  40  per  cent.  Trojan  powder  is  used);  but  more 
40  per  cent,  dynamite  is  used  than  other  kinds  of  explosive.  The 
blasting  is  done  with  electricity,  and  three  XXXXX  detonators  are 
placed  in  each  hole.  The  placing  of  these  holes  requires  considerable 
experience  and  judgment;  so  the  nature  of  the  face  is  studied  and  ex- 
amined for  slips  before  being  drilled.  From  1500  to  3000  cu.  yd.  are 
moved  at  a  blast,  and  about  1500  to  3000  tons  of  ore  when  blasting  in 
ore.  A  churn  drill  will  sink  from  40  to  70  ft.  of  6-in.  hole  in  a  12- 
hour  shift  in  preparing  the  bank  for  blasting,  and  at  times  as  high  as 
120  ft.  has  been  drilled  in  a  shift.  The  boulders  are  bull-dozed.  Two 
No.  5  Keystone  churn  drills  work  on  capping,  and  only  one  on  ore. 

Standard-gauge  equipment  is  used  in  steam-shoveling.  The  capping 
is  loaded  into  dump  cars  of  the  Oliver  type.  Two  sizes  are  used — a 
6-cu.  yd.  and  a  12-cu.  yd.  car  according  to  conditions.  The  larger 
cars  are  equipped  with  standard  air-brake  apparatus,  while  the  6-cu.  yd. 
cars  are  not.  The  larger  cars  are  therefore  better  for  long  runs  and 
require  less  " spotting"  while  being  loaded;  the  small  cars  are  better 
on' curves  and  where  much  dumping  on  trestle  is  done.  The  large  cars 
are  handled  in  trains  of  five,  and  the  small  ones  in  trains  of  eight.  The 
cars  by  actual  measurement  hold  5.4  and  10.9  cu.  yd.,  respectively. 

Bucyrus  shovels  are  used  entirely.  There  are  three  95-ton  shovels 
with  5-cu.  yd.  dippers;  one  95-ton,  with  3  1/2-cu.  yd.  dipper;  and  one 
70-ton  shovel  with  3-cu.  yd.  dipper.  Two  of  these  shovels  are  working 
on  ore,  and  the  rest  on  capping.  About  10  min.  are  required  to  load  a 
50-ton  car  when  running  regular,  but  numerous  delays  increase  this  to  a 
much  lower  average.  The  ore  is  run  directly  to  the  mill  in  these  same  cars, 
but  in  time  some  way  of  screening  the  ore  before  the  mill  bins  are  reached 
will  have  to  be  arranged.  As  it  is,  the  mining  expense  is  increased  con- 
siderably owing  to  the  want  of  storage  capacity  at  the  mine  and  mill. 

The  tracks  are  laid  at  from  3  per  cent,  to  4  per  cent,  grade  and  in 
mining  the  ore  4  miles  of  track  are  used.  Over  this  single-track 


82  MINING    WITHOUT   TIMBER 

system  220  to  230  trains  a  day  are  run.  The  ore-train  yard  is  at  the 
mouth  of  the  pit,  but  the  capping  has  to  be  hauled  some  distance. 
Formerly  this  was  nearly  1  1/2  miles,  but  at  present  the  dumps  are 
much  nearer  and  one  is  within  a  quarter  of  a  mile  of  the  pit. 

In  steam-shovel  operations  the  company  used  six  16x24-in.,  saddle- 
tank  American  Locomotive  Works  locomotives  and  one  14x22-in., 
saddle-tank  locomotive  of  the  same  make.  These  locomotives  use  about 
3  tons  of  coal  in  nine  hours.  Locomotive  engineers  are  paid  $4.25 
for  9  hours;  firemen,  $3  for  9  hours;  switchmen,  $3.25  for  9  hours. 
Trackmen  and  common  labor  is  paid  $2  for  9  hours.  The  powder  boss 
gets  $5  for  9  hours.  Shovel  and  churn-drill  crews  are  alike  in  number 
and  cost  to  those  of  Example  3. 

About  5000  tons  of  ore  and  about  3000  cu.  yd.  of  capping  are  being 
moved  a  day,  or  150,000  tons  of  ore  and  90,000  cu.  yd.  of  capping  per 
month.  On  the  pay  roll  there  are  267  men,  including  men  in  the 
machine  shop,  repair  men,  in  short  every  one  connected  with  steam- 
shovel  mining.  At  present  about  12  acres  of  ground  have  been  stripped, 
but  this  particular  ore  body  has  an  area  of  18  acres. 

In  1910  the  Company's  average  cost  of  shoveling  ore  was  15.4  cents 
per  ton  and  of  removing  waste  was  40 . 6  cents  per  cu.  yd.  Of  the  strip- 
ping cost  15  cents  was  apportioned  to  each  ton  of  ore  extracted  so  that 
the  total  cost  of  mining  the  ore  was  30.4  cents  per  ton  including  re- 
pairs, maintenance,  and  general  expenses. 

EXAMPLE  5. — EASTERN  PENNSYLVANIA  AND  ILLINOIS 
(See  also  Example  49,  51  and  59.) 

Clam-shell  Cranes  and  Wheeled  Dipper-dredges  on  Coal  Seams  with 
Thin  Mantles. — There  are  a  number  of  places  in  the  anthracite  fields  of 
Eastern  Pennsylvania  where  flat  coal  beds  outcrop  so  near  the  surface 
that  they  may  be  stripped  for  100  feet  before  the  cover  attains  an  un- 
profitable thickness.  The  Hilldale  strippings  have  a  section  from  the 
surface  down  about  as  follows: 

Soil,  4.5  ft.;  coal  A,  4.5/J;  rock  3  ft.;  coal  B,  I  ft.;  rock  5  ft.;  coal  C} 
2ft.;  rock  3  ft.;  coal  D,  12  ft. 

Coal  bed  A  has  been  exposed  so  long  it  is  worthless,  except  where  it 
has  a  rock  cover.  Bed  B  is  a  somewhat  rusty  good  coal.  Bed  C  is 
almost  2  ft.  thick  and  could  not  be  worked  at  a  profit  underground;  in 
the  stripping  operations,  however,  it  is,  like  bed  B,  a  source  of  income. 
Bed  D  is  the  coal  aimed  for,  and  is  as  good  as  coal  can  be,  although 
carrying  a  slate  parting. 

The  method  of  stripping  followed  by  Mr.  Kinsley,  the  contractor,  is 
as  follows:  First,  the  top  soil  and  poor  coal  A  is  removed  from  the  top 
rock  by  means  of  a  clam-shell  bucket  and  locomotive  crane.  The  boom 


SURFACE  SHOVELING  IN  OPEN  CUTS  83 

on  this  crane  is  42  ft.  long  and  can  place  the  top  material  where  it  will 
be  out  of  the  way  once  for  all.  The  dirt  is  first  removed  ahead  of  the 
crane  in  the  direction  it  is  moving;  next  from  the  side  of  the  crane  where 
the  mining  is  to  be  carried  on.  The  rock  and  coal  are  then  broken  down 
to  the  coal  bed  D  by  blasting.  The  coal  from  B  and  C  is  picked  out  by 
hand,  while  the  rock  is  wasted  and  piled  back  by  the  bucket  to  form  the 
track  for  the  coal  car.  The  coal  bed  D  is  next  broken  by  powder  and 
loaded  into  the  bucket  by  hand.  The  bucket  is  then  raised  by  the  crane 
and  the  coal  dumped  in  the  car.  Anthracite  coal  is  so  brittle  it  breaks 
in  handling,  and  while  it  was  at  first  intended  to  use  the  bucket  to  pick 
up  the  coal,  it  was  found  inadvisable  to  do  so.  This  method  of  stripping 
is  equivalent  to  making  a  side  cut  along  the  crop  for  the  crane  track, 
then. excavating  below  this  cut,  and  filling  in.  The  fill  will  furnish  the 
road  for  the  next  side  cut  when  coming  back  from  the  boundary  line  of 
the  property,  as  it  does  for  the  coal  cars  in  going  forward. 

Near  Danville,  111.,  there  was  a  35-acre  bituminous  coal  bed,  8  ft. 
thick,  which  was  approximately  horizontal.  This  coal  bed  outcropped 
on  all  sides  of  a  flat-topped  hill,  and  had  an  overburden  from  38  to  40  ft. 
thick,  composed  of  soil,  clay,  gravel,  and  about  20  ft.  of  shale.  It  was 
decided  to  strip  to  the  coal  with  a  single  cut  and  to  take  a  wide  cut  so  as 
better  to  provide  for  the  efficient  mining  of  the  coal.  The  plant  decided 
on  was  a  steam  shovel,  made  by  Belief ontaine  (Ohio)  Machine  Co., 
having  a  dipper  of  2  cu.  yd.  capacity,  and  mounted  on  a  movable 
platform,  provided  with  a  Jeffrey  belt  conveyer  for  disposing  of  the 
material.  The  platform,  which  was  30  ft.  wide,  was  mounted  on  four 
trucks  that  were  moved  as  desired  on  two  tracks.  The  machine,  as  it 
appeared  in  operation,  is  shown  in  the  frontispiece.  After  excavating 
the  material  above  the  coal  with  the  bucket,  it  was  swung  to  a  large  steel 
hopper  and  discharged.  From  the  bottom  of  the  hopper  the  material 
was  carried  on  a  steel  cross-feeder  to  the  lower  end  of  the  belt  conveyer. 
The  latter  was  a  40-in.  wide  belt  traveling  on  a  steel  arm  105  ffc.  long. 
The  arm  was  supported  by  wire  ropes  from  a  tower  built  above  the 
platfoim  to  a  height  <"f  48  ft.  By  this  arrangement  the  waste  clearance 
at  the  outer  end  of  the  arm  was  about  60  ft.  above  the  tracks. 

The  machine  is  said  to  have  had  no  difficulty  in  excavating  heavy 
pieces,  stumps,  and  logs,  and  depositing  them  in  the  space  where  the  coaJ 
had  been  mined.  After  the  overburden  had  been  removed  the  coal  was 
quarried  and  loaded  into  cais  on  a  track  laid  in  the  place  from  which 
coal  had  been  taken  pieviously.  As  the  shovel  made  a  cut,  the  debris 
was  deposited  in  the  cavity  made  by  removing  the  coal.  The  tracks  for 
the  machine  we^e  laid  on  top  of  the  coal,  which  was  mined  so  as  to  leave 
a  bench  for  the  shovel  to  come  back  on. 


CHAPTER  VII 


SURFACE  MINING 

EXAMPLE  6— PUERTOCITOS  MINE,  CANANEA,  SONORA,  MEXICO 
(See  also  Examples  IS,  34  and  45.) 

Quarrying  Sidehill  Lenses  with  Rock  Capping. — The  ore  deposit  is 
on  a  sidehill  (see  Fig.  35)  and  consists  of  streaks  of  malachite  through  a 
bedded  limestone.  On  the  surface  the  metallic  contents  have  been 
leached  so  that  a  worthless  limestone  capping  must  first  be  removed 
before  attacking  the  ore  of  which  two-thirds  can  be  rejected  during 
sorting.  The  capping  and  waste  are  dumped  from  small  cars  down 


ft: 


FIG.  35. — Openpit  mining  at  Puertocitos. 

chutes  into  pockets  above  the  railroad  whence  it  can  be  hauled  to 
nearby  dumps  or  used  as  filling.  The  ore  is  dumped  into  similar  pockets 
and  is  then  hauled  10  miles  downhill  to  the  smelter  for  about  12  cents 
per  ton. 

In  removing  the  capping,  drill  holes  are  put  down  only  to  its  bottom, 
but  in  the  ore  regular  benches  are  laid  off,  22  ft.  high.  The  drilling  is 
by  hand  with  three  men  to  a  crew.  Down  to  a  depth  of  9  ft.  the  hole  is 
drilled  by  hammers,  there  being  two  hammermen.  This  takes  about 
half  a  day.  As  that  is  the  limit  of  economic  work  with  hammers,  from 

84 


SURFACE    MINING  85 

that  point  the  hole  is  churned  down  by  the  three  men  to  a  depth  of  22  ft. 
This  takes  a  day  and  a  half  more,  or  two  days  to  complete  a  hole  22  ft. 
deep.  These  holes  are  sprung  with  dynamite,  and  then  loaded  with 
from  four  to  six  kegs,  50  Ib.  each,  of  black  powder,  since  the  holes  are 
generally  drilled  with  a  toe  of  ground  at  the  bottom  about  equal  to 
the  depth.  The  holes  are  also  spaced  about  equally,  although  the  manner 
of  breaking  of  the  holes  previously  blasted  determines  the  position  of 
the  next  holes.  Each  blast  breaks,  on  an  average,  about  600  tons  of 
rock.  Electric  blasting  is  used  in  firing  these  holes,  and  several  are 
blasted  at  a  time. 

The  tramming  distances  are  short,  but  owing  to  the  large  amount  of 
sorting  necessary,  the  block-holing  and  the  sledging  required  in  breaking 
up  the  pieces  for  sorting,  the  cost  of  placing  ore  on  the  cars  is  about  $1 . 70 
a  ton,  including  stripping  and  all  other  charges.  Still,  if  much  powder 
were  used  so  as  to  avoid  the  sledging,  the  rock  would  be  broken  so  fine 
that  much  ore  would  be  lost  that  might  otherwise  be  sorted  out,  and 
the  fines  might  become  too  low  grade  to  be  sent  to  the  smelter. 

The  mine  is  sending  to  the  smelter  an  average  of  140  tons  of  ore  per 
day,  and  employs  about  175  men.  These  are  all  Chinamen,  as  it  is 
hard  to  get  Mexicans  to  work  outside  during  the  winter  months,  for  at 
this  altitude  the  climate  is  rigorous  and  there  is  often  snow. 

EXAMPLE  7. — MESABI  IRON  RANGE,  MINN 
(See  also  Examples  2  and  46.) 

Milling  or  "Glory-hole"  System  for  Flat  Lenses  under  Glacial  Drift 
(Vertical  Chutes). — The  milling  system,  having  little  machinery,  costs 
much  less  to  start  than  open-cutting  by  steam  shovels,  but  the  cost  of 
ore  per  ton  in  the  extraction  is  greater  on  account  of  underground  de- 
velopment, tramming,  hoisting  and  lighting.  It  is  adapted  to  deposits 
where  the  overlying  mantle  is  not  so  deep  as  to  necessitate  underground 
mining  and  where  facilities  of  approach  and  the  general  shape  of  the 
ore  body  are  unsuitable  for  open-cutting.  Like  the  latter,  milling 
cannot  be  worked  so  cheaply  in  winter  and  is,  therefore,  best  adapted  to 
deposits  which  need  only  make  shipments  during  the  open  season  on 
the  lakes.  Milling  can  often  be  used  as  an  auxiliary  to  open-cutting  to 
remove  such  parts  of  the  stripped  ore  body  which  are  so  deep  or  so 
obstructed  as  not  to  allow  of  easy  railroad  grades  from  the  pit's  approach. 
Many  mines  first  opened  on  a  small  scale  by  milling  have  since  been 
extended  for  open-cutting  by  steam-shovels. 

In  starting  a  mine  on  this  system  it  is  only  necessary  to  strip  suffi- 
cient area  so  that  the  top  diameter  of  the  final  funnel-like  ore  pit  will  be 
adequate  to  allow  the  base  of  the  ore  to  be  reached  by  a  series  of  con- 
centric, descending  benches  of  economical  height  and  width.  While 


86 


MINING    WITHOUT   TIMBER 


stripping,  a  working  shaft  s  (Fig.  36),  with  two  skip  compartments 
and  one  cage  compartment,  is  sunk  in  the  wall  rock  alongside  the  ore 
and  a  crosscut  c  run  from  the  foot  wall  to  a  point  under  the  proposed 
mill,  where  a  vertical  raise  r  is  made  to  the  surface.  By  extending  a 
drift  d  from  the  foot  of  this  raise  along  the  ore  body  and  raising  from 
it  at  appropriate  intervals  (50  ft.  to  100  ft.)  as  many  mill  holes  can  be 
started  as  the  area  stripped  will  permit. 

The  stoping  begins  at  the  top  of  of  a  mill  by  circling  it  with  an  under- 
hand bench  b.  The  height  of  a  bench  is  governed  by  the  depth  of 
the  hole  that  can  conveniently  be  bored  with  the  piston  air  drill  used; 
usually  between  10  ft.  and  20  ft.  The  width  of  a  bench  depends  upon  the 


FIG.  36. — Milling  or  glory-hole  system. 

economical  burden  that  one  line  of  holes  will  carry  when  chambered  by. 
squibbing,  and  also  upon  the  width  that  can  clear  itself  after  blasting, 
mostly  by  gravity. 

When  the  benches  have  reached  the  bottom  of  the  ore  body  and  the 
funnel  is  completed  along  line  f-q-m-n-g-h-k-t-x-y  (if  there  is  only  one 
mill)  the  broken  ore,  if  the  benches  are  cut  down  much  farther,  will  no 
longer  slide.  Then  the  benches  have  to  be  cleared,  and  for  this  purpose 
a  steam  shovel  can  sometimes  be  efficiently  used.  In  Fig.  27  the  great 
"mill  pit"  was  formed  by  cutting  down  and  uniting  a  number  of  mill 
funnels.  At  the  bottom  of  each  mill  is  a  chute  gate,  from  which  the 
ore  is  drawn  into  cars  to  be  trammed  to  the  shaft,  whence  it  is  hoisted 
in  a  skip  to  the  surface.  As  the  Mesabi  ore  is  friable,  there  is  seldom 
trouble  from  the  clogging  of  chutes  by  boulders. 

In  other  districts  where  steam  shovels  for  stripping  are  not  so  easily 
available  as  on  the  Mesabi  or  where  the  topography  or  repair  facilities 
are  unsuitable  for  their  operation,  the  surface  waste  over  the  ore  body 
can  be  removed  by  several  other  methods. 

The  first  to  suggest  itself  would  be  to  carry  the  raise  of  each  mill- 
hole  up  to  the  surface  and  drop  all  the  waste  through  it  for  disposal  on  a 
dump,  near  the  exit  of  shaft  or  adit,  before  attacking  the  ore  beneath. 


SURFACE    MINING 


87 


This  would  involve  handling  all  the  waste  in  adit  or  shaft  and  often  a 
cheaper  way  of  stripping  would  be  by  such  rapid  methods  as  horse 
scrapers,  drag-line  excavators,  hoist-cableways  or  hydraulicing. 

Sometimes  the  broken  ore  is  hoisted  from  the  bottom  of  the  open 
pit  by  a  clam-shell  or  other  self-filling  bucket  suspended  from  a  hoist- 
cableway  stretched  between  derricks  on  the  surface,  and  as  this  system 
saves  the  cost  of  tunnel  and  shaft  it  is  often  cheaper  for  small  deposits. 

EXAMPLE  8. — TRADERS'  MINE,  MENOMINEE  RANGE,  MICH 
(See  also  Examples  38  and  46.) 

Milling  or  "  Glory -hole"  System  for  Vertical  Wide  Vein  Without 
Mantle  (Vertical  Chutes). — This  ore  body  is  100  ft.  to  200  ft.  thick  and 
has  little  or  no  overburden.  It  extends  conformably  to  the  enclosing 
beds  of  the  Traders'  formation  (named  from  the  mine)  for  over  a  mile, 
and  dips  60  deg.  south,  the  foot  wall  being  of  greenish  and  the  hangwall  of 
reddish  slate.  Though  a  few  high-grade  pockets  have  been  extracted, 
the  bulk  of  the  mines'  output  of  1,500,000 
tons  has  been  low-grade  ore  of  41  per  cent, 
iron  and  0.015  per  cent,  to  0.018  per  cent, 
phosphorus.  The  iron  mineral  is  high  grade, 
but  it  occurs  only  in  bands  in  a  hard  jaspilite 
matrix.  The  annual  output  is  but  125,000 
tons,  as  work  goes  on  only  during  the  season 
of  open  lake  navigation.  The  Antoine  Ore 
Co.  is  the  operator.  For  a  daily  output  of 
900  tons  to  1,000  tons  there  are  125  men,  of 
whom  64  are  machine  men  (16  drills  on  each 
shift),  and  most  of  the  balance  are  trammers. 
The, low  mining  cost  of  30  cents  per  ton  is  less  J 
than  that  for  steam  shoveling  on  the  Mesabi  FIG.  37.— Chute  for  milling  system, 
where  the  stripping  is  heavy. 

Development. — The  open  pit  was  opened  by  a  4-compartment  shaft 
in  the  foot  wall,  whose  sump  is  150  ft.  from  the  collar  and  225  ft.  below 
the  apex  of  the  vein's  outcrop.  It  is  7  ft.  by  23  ft.  inside  of  timbers  and 
has  a  5-ft.  by  7-ft.  cageway,  two  6-ft.  by  7-ft  skipways  and  a  3-ft.  by 
7-ft.  pipeway,  all  dividers  and  wall  plates  being  12  in.  by  12  in.  From 
this  shaft  a  crosscut  runs  on  the  80-ft.  level  to  the  main  drift,  which 
extends  for  1,450  ft.  along  the  center  of  the  vein.  From  the  main  drift 
are  turned  off  crosscuts  at  100-ft.  intervals,  and  where  they  strike  the 
foot  wall  a  vertical  raise  6  ft.  square  is  put  up  to  the  outcrop  (as  in  Fig. 
36)  for  the  beginning  of  a  mill  hole. 

No  timber  is  used  in  drifts,  crosscuts  or  raises  except  for  the  chutes. 
The  bottom  of  the  raise  R  (Fig.  37)  is  rock-filled  and  floored  with  poles 


88  MINING    WITHOUT   TIMBER 

p,  which  are  set  to  as  to  leave  a  loading  orifice  2  ft.  by  5  ft.,  controlled 
by  a  steel  bar  6  resting  in  cramps  c.  Two  drift  sets,  S,  are  inserted  oppo- 
site the  raise  to  complete  the  chute  with  8-ft.  caps  and  posts. 

For  each  mill  hole  there  are  two  31/4  in.  Rand  drills  on  tripods,  which 
bore  down  holes  8  ft.  to  14  ft.  long  and  are  supplied  with  air  by  small  pipes 
from  the  top  of  the  pit.  As  the  ore  is  broken  in  big  chunks  these  have  to 
be  reduced  to  prevent  choking  the  chutes,  and  this  is  done  by  bull-dozing 
rather  than  block-holing,  as  the  extra  powder  is  found  less  expensive 
than  the  extra  labor  for  boring  block  holes.  Either  the  chute-raise 
must  be  kept  full  of  ore  or  the  entrance  of  each  chute  at  the  base  of  the 
underhand  benches  must  be  covered  over  by  a  log-grizzly  to  arrest  the 
descent  of  the  boulders  into  the  chute  until  they  can  thus  be  broken.  The 
tram  cars  t  are  handled  by  two  men  and  hold  two  tons,  or  one  skipload. 

EXAMPLE  9. — ALASKA  TREADWELL  GOLD  MINE,  DOUGLAS  ISLAND, 

ALASKA 

(See  also  Example  16.) 

Milling  or  "Glory-hole"  System  for  Sub-vertical  Wide  Veins  Without 
Mantle  (Hour-glass  Chutes). — The  orebodies  her  occupy  a  huge  syenite 
dike  that  has  intruded  the  slate  country  rock  for  several  miles.  The 
dike  is  irregular  in  width,  varying  from  420  ft.  at  the  Treadwell  mine  to 
150  ft.  at  the  Mexican  and  300  ft.  at  the  Ready  Bullion,  a  half  mile  to  the 
southeast,  while  in  the  interval  between  these  three  miles  the  dike  is  a 
mere  stringer.  The  mineralization  of  the  syenite  was  due  to  a  subse- 
quent intrusion  of  barren  gabbro  which  now  forms  the  hangwall  of  the 
syenite  orebodies  and  in  places  is  badly  schattered.  There  is  also  a  third 
intrusion  of  barren  basalt  which  is -found  in  the  orebodies  as  a  single 
dike  above,  but  as  several  smaller  ones  at  depth. 

The  chief  ore  is  of  two  varieties:  first,  stringers  of  quartz  and  calcite 
occupying  fracture  planes  in  the  syenite;  second,  crushed  and  broken 
syenite  which  has  been  saturated  by  mineral-bearing  solutions.  The 
largest  orebody  is  the  Treadwell  shown  in  section  in  Fig.  38,  which  dips 
70°  and  has  a  gabbro  (greenstone)  hangwall  and  a  black  slate  footwall. 
The  climate  is  wet  but  mild  enough  to  permit  continuous  outdoor  mining, 
so  that  the  main  openpit  finally  reached  a  depth  of  220  ft.  below  the  adit 
level  and  450  ft.  from  the  surface  with  a  maximum  width  of  420  ft.  and 
a  length  of  1700  ft.  The  large  slides  of  waste  rock  from  the  footwall 
and  the  need  of  a  thick  pillar  of  rock  to  protect  the  underground  workings 
from  surface  water  caused  the  stoppage  of  the  openpit  at  the  220-ft. 
level. 

To  develop  the  Treadwell  mine  below  the  Adit  Level,  a  four-com- 
partmant  vertical  shaft  (sse  Fig.  38)  was  sunk  in  the  hangwall  and  stations 
as  wide  as  the  shaft  and  40  to  60  ft.  long  were  cut  at  each  level.  A 


SURFACE    MINING 


89 


main  crosscut,  as  C,  is  run  on  each  level  to  the  footwall,  20  ft.  wide  for 
the  first  100  ft.  and  12  ft.  in  width  for  the  balance  of  the  distance.  Be- 
neath the  floor  of  the  station  an  ore  bin,  B,  is  cut  out  in  the  rock  with  a 
capacity  of  500  to  1500  tons,  to  afford  ample  storage.  On  the  main 
crosscut,  on  its  hangwall  end,  is  cut  a  station  for  the  winding  engines  for 
the  tail  rope  system  of  haulage.  Directly  opposite  the  sinking  com- 
partment, on  alternate  levels,  a  station  is  cut  for  the  sinking  hoist. 


///  LEGEND 

#Solid  Ore 
Broken  Ore 


FIG.  38. — Alaska  Ireadweil  mine,  crods  section. 

When  the  main  crosscut  has  reached  the  footwall,  parallel  drifts,  D, 
are  turned  off,  at  right  angles  and  about  60  ft.  apart,  to  follow  the  strike 
of  the  vein.  At  intervals  of  25  ft.,  raises  are  now  put  up  on  alternate 
sides  of  both  crosscut  C  and  drifts  D.  These  raises  are  15  ft.  high, 
have  a  slope  of  60  deg.,  so  that  the  ore  will  run  freely,  and  are  fitted 
with  special  finger-chutes  so  the  large  quantities  of  ore  can  be  easily 
run  into  the  mine  cars.  At  the  same  time  as  these  drifts  and  raises, 
there  are  being  run  intermediate  drifts  (as  E)  directly  above  each  drift 
D  but  separated  from  its  back  by  a  rock  pillar,  10  ft.  thick. 

When  a  pit  P  is  to  be  opened,  a  raise  is  put  up  from  the  nearest  level 
and  connected  with  the  surface.  This  raise  is  started  from  an  intermedi- 
ate drift  E,  in  general  directly  over  a  chute-raise.  The  chutes,  25  ft. 


90  MINING    WITHOUT    TIMBER 

distant  on  each  side,  then  serve  as  man-ways  for  the  raise  in  course  of 
erection,  and  the  broken  rock  is  drawn  off  through  the  middle  chute- 
raise  into  cars.  When  the  raise  has  been  connected,  the  machine-drills 
are  put  to  work  cutting  out  a  small  stope  of  the  bottom.  Thus  the 
raise  when  finished  has  the  shape  of  an  hour-glass,  the  top  being  formed 
by  the  open  pit  P  and  the  bottom  by  a  drawing-off  stope  G,  covering 
three  chutes  and  from  20  to  30  ft.  high,  the  two  being  joined  by  the 
raise.  The  object  of  cutting  out  the  pit-raises  in  this  manner  is,  first, 
to  obtain  chute-capacity  in  case  of  their  being  hung  up  by  large  pieces 
of  rock  or  by  blasting;  and,  second,  to  afford  an  opportunity  to  break 
up  any  large  piece  of  rock  that  may  have  been  overlooked  in  the  pit, 
which  would  stop  up  the  chute  unless  it  were  broken  to  pieces  small 
enough  to  pass  through  it. 

Machine-drilling  is  seen  at  its  best  in  these  pits.  The  3  1/4-in.  di- 
ameter Ingersoll-Sergeant  drills,  set  on  tripods,  are  used  in' all  the  pits  at 
present.  The  average  number  of  feet  drilled  per  machine  in  10  hours  is 
36.35.  The  holes  are  drilled  to  an  average  depth  of  12  ft.,  and  each 
machine  will  break  69.69  tons  of  ore  per  shift  of -10  hours.  When  the 
pits  were  smaller  and  the  difficulty  of  setting  up  was  not  so  great  as  at 
present,  the  average  number  of  feet  drilled  was  much  higher,  and  the 
breaking  capacity  of  a  machine-drill  was  from  150  to  200  tons  of  ore 
shift  of  10  hours.  The  pits  are  worked  by  drilling  and  blasting  the  ore 
from  a  series  of  benches  or  terraces  around  the  chute-raise  as  a  center, 
and  when  the  ore  is  blasted  the  broken  rock  rolls  down  to  the  bottom. 
The  small  pieces  are  then  broken  by  sledges,  and  the  larger  ones  by 
placing  sticks  of  powder  on  the  surface  of  the  rock,  tamping  with  a  little 
fine  dirt,  and  blasting.  For  blasting  holes,  No.  2,  or  40  per  cent.,  dyna- 
mite is  used,  while  for  "bulldozing"  No.  1,  or  70  per  cent.,  is  best. 

When  the  rock  has  been  broken  to  the  required  size,  it  is  drawn  off, 
through  the  raises  and  chutes  described  above,  into  cars.  These  cars 
are  hauled  to  the  station  ore-bins  by  horses,  or  by  endless-rope  haulage, 
where  they  are  dumped.  The  ore  is  then  loaded  into  skips  and  hoisted 
to  the  surface. 


CHAPTER  VIII 
UNDERHAND  STOPING 

EXAMPLE  10. — DISSEMINATED  LEAD  FIELD  OF  SOUTHEAST  MISSOURI 

Underground  Quarrying  with  Down  Holes  in  Flat  Lenses  in  Limestone. — 
Topographically  the  country  is  hilly,  even  rugged  in  places,  and  is 
traversed  by  a  network  of  small  streams.  The  elevation  varies  from 
500  to  1000  feet  above  the  sea  level.  The  surface  is  well  wooded  with 
forests  of  oak,  hickory,  ash,  and  yellow  pine,  the  first  predominating. 

The  surface  formation  is  of  Cambrian  sediments  which  abut  against 
hills  of  Archean  granite.  The  Cambrian  consists  here  of  the  St.  Francois 
limestone,  which  in  places  is  700  ft.  thick,  resting  conformably  on  the 
Mine  LaMotte  Sandstone.  The  ore  bodies  lie  entirely  within  this  limestone, 
and  its  base  represents  the  lowest  points  to  which  shafts  have  been  sunk; 
the  shaft  depths  varying  from  90  ft.  at  Doe  Run  and  Mine  La  Motte, 
to  600  ft.  at  Flat  River.  The  greatest  proved  run  of  pay  ore  is  that  at 
Bonne  Terre,  whose  length  is  over  half  a  mile,  width  up  to  200  ft.,  and 
height  25  to  100  feet. 

Prospecting. — The  great  development  since  1890  has  been  entirely 
based  upon  results  obtained  by  systematic  prospecting  with  the  diamond 
drill,  which  was  introduced  at  Bonne  Terre  by  the  St.  Joe  Lead  Co.,  in  1869. 
The  lead  areas,  underlying  the  country  horizontally  under  shallow  depths 
of  homogeneous  limestone,  make  conditions  unusually  favorable  for  this 
form  of  prospecting.  The  gently  undulating  topography  offers  no  diffi- 
culty to  the  movement  of  the  drills,  while  the  warm  climate  permits 
out-of-door  work  for  most  of  the  year.  A  portable  drilling  outfit  is  used, 
with  the  drill  and  pump  attached  to  a  "agricultural"  boiler  on  wheels. 

The  distance  drilled  in  a  shift  is  very  large  compared  with  vertical 
vein  practice.  Sixty  feet  of  uncored  hole  in  10  hours  is  a  common  aver- 
age for  holes  500  ft.  deep,  and  as  much  as  100  ft.  is  often  run.  The 
coring  is  not  begun  until  the  lead-bearing  zone  is  approached,  and  then 
20  to  30  ft.  per  shift  is  commonly  done.  The  bit  used  is  set  to  cut  a  hole 
21/8  in.  diameter,  with  usually  four  diamonds  outside  and  four  inside  of 
the  bit.  The  filler,  with  inside  diameter  of  1/2  in.,  to  make  the  bit 
solid,  is  set  with  four  diamonds  more  of  two  to  three  carats  apiece, 
rounded  carbons  being  prefered. 

As  much  of  the  sludge  never  reaches  the  surface,  being  filtered  off 
through  cracks,  sampling  it  would  be  of  no  value;  so  the  core  is  entirely 
relied  on,  both  for  locating  the  deposits  vertically  and  for  giving  the 

91 


92  MINING    WITHOUT    TIMBER 

assay  yield.  The  sludge,  however,  gives  the  drill  man  warning  of  the 
approach  of  the  lead  horizon.  The  cost  of  drilling  500  ft.,  with  dia- 
monds at  $40  a  carat,  is  not  far  from  $0 . 50  per  foot  at  Flat  River. 

In  locating  the  first  hole  in  a  virgin  tract  an  endeavor  is  made  to 
locate  the  extension  of  the  axis  of  an  adjoining  lead  run.  If  there  are  no 
nearby  ore  runs,  surface  crevices  and  ancient  diggings  are  looked  for  as  a 
guide  to  the  occurrence  of  a  possible  ore  body  underneath.  If  neither 
adjoining  runs  or  surface  indications  are  present,  the  only  recourse  is  to 
lay  the  area  off  in  5,  10,  or  20  acre  squares,  according  to  the  probability 
of  finding  ore,  and  bore  a  hole  in  the  center  of  each  square.  If  one  hole 
strikes  an  ore  body,  a  circle  of  209  ft.  radius  is  struck  off,  and  holes  bored 
on  this  circumference  209  ft.  apart  to  determine  the  axis  of  the  lead  run. 

For  traversing  the  surface  broken  ground  which  is  sometimes  100  ft. 
deep,  the  St.  Louis  Prospecting  Co.  used  a  portable  steam  churn-drill  outfit 
in  the  following  way :  To  penetrate  the  surface  soil,  the  bit's  cutting  edge 
is  made  7  inches.  This  allows  a  casing  of  5  5/8  in.  inside  diameter  to 
be  used  that  is  inserted  2  ft.  into  the  bed  rock.  The  drilling  is  then 
pushed  downward  until  a  rock  suitable  for  diamond  drilling  is  encount- 
ered, with  a  bit  of  51/2  in.  diameter.  This  diameter  gives  room  for  the 
placing  of  an  inserted  joint  casing  of  5  in.  inside  and  53/8  in.  outside 
diameter  at  the  upper  end,  in  case  mud  channels  or  opinings  are  struck 
so  large  as  to  make  progress  without  the  casing  impossible;  a  speed  of 
10  ft.  in  11  hours  was  the  average  made  in  this  kind  of  work. 

Development. — The  more  thoroughly  the  ore  has  been  drilled  the 
easier  it  is  to  lay  out  the  shaft  and  drifts  advantageously.  The  problem 
is  similar  to  that  of  a  coal  seam  development,  but  unlike  the  usual 
bituminous  seam,  the  lead  run  is  irregular  both  in  thickness  and  in  the 
level  of  its  floor.  The  shaft  is  located  from  three  considerations:  1,. 
the  lowest  point  of  the  ore  body;  2,  the  center  of  the  pay  ore  body; 
3,  the  propinquity  to  a  good  mill  site. 

The  customary  way  of  horizontal  advance  is  breast  stoping,  but 
drifting  is  used  to  reach  pockets  separated  from  the  main  run  of  ore,  or 
to  traverse  barren  places.  To  reach  the  top  of  the  ore  body  to  start  a 
stope,  either  vertical  or  inclined  raises  are  used.  Fig.  39  shows  the  method 
of  vertical  raising.  As  soon  as  enough  is  excavated  above  the  floor  the 
two  stulls  are  put  in,  a  floor  of  poles  laid  on  them,  and  the  raise  then 
continued  in  two  compartments  to  the  top,  the  chute  being  separated 
from  the  ladder-way  by  a  pole  partition.  The  broken  rock  is  left  in  the 
chute  for  the  men  to  stand  on,  the  drill  bar  is  wedged  horizontally 
against  the  walls  and  the  surplus  broken  rock  is  thrown  down  the  ladder- 
way.  When  the  raise  is  fin'shed,  the  floor  is  blasted  out  and  the  rock 
loaded.  A  gate  is  only  put  in  the  chute  at  the  bottom,  if  the  chute  is  to 
be  used  for  the  lowering  of  rock  after  the  completion  of  the  raise,  which 
often  occurs  in  exhausting  ore  bodies  above  the  main  level  in  operation. 


UNDERHAND    STOPING 


93 


The  inclined  raises  are  run  the  same  as  drifts  and  as  steep  (45  deg.) 
as  's  possible  without  using  timbers  to  hold  up  the  broken  rock  that 
the  men  stand  upon.  Whether  the  vertical  or  the  inclined  raise  is  used 
to  open  up  a  stope  depends  on  the  shape  of  the  ore  body. 

Sloping. — There  is  no  timbering  in  the  drifts  or  stopes,  except 
sprags  for  supporting  piping,  and  no  filling  system.  The  roof  is  held  up 
by  pillars  which,  however,  can  be  laid  out  on  no  regular  plan  as  in  coal 
mining.  As  far  as  possible,  pillars  consist  of  lean  ore.  The  width  of 
the  stope  between  pillar  depends  on  the  strength  of  the  roof  and  varies 


Stpk 


Drift 


Long.  Sect. 


Plan 
FIG.  39. — Raise,  S.  E.  Mo. 


FIG.  40.— Stope,  S.  E.  Mo. 


from  40  ft.  at  the  Desloge  and  Bonne  Terre  mines,  to  15  or  20  ft.  at  the 
Theodora,  and  other  eastern  Flat  River  mines. 

The  underhand  stoping  system  is  usod.  A  breast  6  ft.  high  is  run 
from  the  top  of  raise  along  the  roof  of  the  ore  body,  which  is  tested  for 
ore  above  by  driving  an  upper  air-drill  bole  into  it  occasionally,  and  as 
many  benches  quarried  out  beneath  it  as  are  necessary  to  excavate  the 
ore  to  the  level  below. 

There  are  two  systems  of  breast  stoping,  one  for  narrow  and  the 
other  for  the  wide  s'opes.  In  Fig.  40  a  breast  15  ft.  wide  is  shown, 
the  holes  are  6  to  8  ft.  deep  and  are  nine  or  ten  in  number  in  three  or 
four  vertical  and  horizontal  rows.  Each  row  is  fired  in  order  1,  2,  3  and 
4,  and  causes  an  advance  of  but  2  ft.  in  tha  breast,  each  side-cut  being 
kept  on  the  same  side  of  the  breast  to  turn  off  a  round  pillar.  Two 


94 


MINING    WITHOUT    TIMBER 


men  with  a  machine  can  drill  the  blast  a  round  of  nine  holes  in  a  10  hour 
shift,  the  drill  bar  being  set  up  once  for  each  vertical  row  of  holes. 

For  the  wider  stopes  a  similar  method  is  used;  but  an  advance  of 
3  to  5  ft.  instead  of  2  ft.  is  scored  for  each  round,  the  holes  for  this 
being  placed  nearer  across  the  breast,  the  rows  A,  B,  C,  etc.,  Fig.  40, 
breaking  from  2  to  4  ft.  each.  The  machine  men  drill  complete  as  many 
vertical  rows  as  possible  and  blast  them  before  going  off  shift. 

\Yhen  the  face  of  the  breast  stope  is  10  to  12  ft.  from  the  side  of  the 
raise,  a  bench  stope  is  started.  The  drills  for  this  are  set  on  a  tripod 
and  the  holes  are  6  to  10  ft  deep.  The  holes  take  off  2  to  4  ft.  of  burden; 
and  they  are  placed  from  4  to  8  ft.  apart  along  the  bench,  the  center 
one  being  fired  first  and  then  the  sides.  The  bench  holes  are  farther 
apart  in  the  wide  stopes. 


FIG.  41.— Ladder  scaffold,  Ducktown,  Tenn. 

In  firing  a  stope  the  lower  bench  is  fired  and  the  bench  holes  the  last 
of  all;  the  holes  being  ignited  simultaneously  at  the  end  of  the  shift. 
The  fuses  are  cut  of  graded  lengths.  One  drill  can  bores  six  to  eight 
bench  holes  in  a  shift,  and  in  the  15  ft.  stope  of  Fig.  39,  two  bench 
drills  can  keep  pace  with  three  breast  machines — which  ratio  does  not 
alter  appreciably  for  wider  stopes,  for  breast  as  well  as  bench  stoping 
gains  in  speed  in  them.  A  pound  of  powder  will  break  three  times  as 
much  rock  on  a  bench  as  in  a  breast  stope.  One  air  drill  in  a  Flat  River 
producing  mine  with  stopes  12  to  20  ft.  high  will  drill  enough  to  break 
20  to  30  tons  of  rock  in  two  shifts,  while  at  Bonne  Terre  in  stopes  20 


UNDERHAND    STOPING  95 

to  60  ft.  high,  one  drill  will  break  over  40  tons  in  the  same  time. 
The  explosive  for  stopes  is  35  per  cent,  dynamite  and  1  Ib.  will  break 
from  1  ton  of  rock,  in  a  narrow  breast,  to  4  tons  on  a  wide  and  deep 
bench. 

Though  the  roof  is  self-sustaining,  pieces  are  liable  to  shell  off  BO 
that  a  roof  man  is  necessary  to  bar  down  the  loose  pieces  with  gad  and 
pick;  the  sides  of  the  shaft  have  also  to  be  occasionally  inspected  and 
barred  down.  The  waste  Lmestone  is  broken  down,  loaded,  and  put 
on  the  surface  dump  as  it  seldom  convenient  to  store  it  underground. 

Ladder  Scaffold,  Ducktown,  Tennessee. — In  the  similar  system  of 
underhand  stoping  in  the  Tennessee  Copper  Company's  mines,  where 
the  miners  practically  never  see  the  back,  which  in  an  open  stope 
is  frequently  from  70  to  90  ft.  above  them,  it  is  evidently  necessary 
to  keep  the  roof  well  trimmed  of  all  heavy,  or  "balk  ground." 
To  insure  this,  a  crew  of  men  is  continually  kept  at  work,  looking 
after  the  condition  of  the  roof.  This  work  is  extremely  dangerous 
and  ready  resource  is  required  to  enable  the  men  to  gain  access  to 
the  back.  Fig.  41  shows  the  method  of  rigging  ladders  to  reach  the 
roof  over  the  benches  of  an  underhand  stope,  open  to  its  full  height 
and  for  a  width  of  from  50  to  150  ft.  The  ladders  are  securely  lashed  to- 
gether, and,  as  shown,  stayed  by  ropes  secured  to  the  drill  steels  set 
into  the  rock  face.  A  small  stoping  drill  is  frequently  slung  from  the 
ladder  and  used  to  put  holes  in  the  roof  where  much  balk  ground  must 
be  slabbed  down.  Shooting  the  roof,  is,  however,  a  dangerous  practice, 
as  shattered  rock  is  apt  to  be  left  to  fall  later,  when  the  face  of  the  stope 
has  advanced  and  the  back  is  inaccessible. 

EXAMPLE   11. — DAVEY  MINES,  AMERICAN  ZINC,  LEAD,  AND  SMELTING 
COMPANY,  JOPLIN,  SOUTHWEST  Mo  * 

Underground  Quarrying  with  Horizontal  Lenses  in  Limestone. — The 
zinc  and  lead  ores  of  this  typical  sheet-ground  formation  vary  from  16 
ft.  to  20  ft.  in  thickness,  and  underlie  continuously  a  large  area.  The 
depth  of  the  deposit  is  about  240  ft.  The  upper  6  ft.  is  considerably 
the  richest  portion  of  the  sheet  or  bed  and  carries  most  of  the  "jack"  or 
-sphalerite.  The  lower  portion  contains  most  of  the  lead,  which  occurs 
in  pockets  or  in  irregular  sheets. 

The  shaft  of  Davey  No.  3  mine  is  256  ft.  deep  and  is  9x18  ft.  all 
the  wray  down.  This  size  is  larger  than  necessary,  the  shafts  later  sunk 
being  7x12  ft.  To  sink  a  7x12  ft.  shaft  in  limestone  costs  about  $20  per 
foot.  No.  3  shaft  has  two  9x5  1/3-ft.  hoisting  compartments,  and  a  pipe 
compartment  at  each  end.  The  timbering  is  very  simple,  consisting  of 
4x9-in.  stulls  set  centered  every  5  ft.  4  in.,  over  which  is  nailed  a  lining  of 
Ixl2-in.  boards. 


96 


MINING    WITHOUT   TIMBER 


The  chert  rock  of  the  ore-bearing  sheet  formation  breaks  quite 
readily  because  of  its  brittleness  and  compactness.  The  system  used  for 
breaking  ground  is  as  follows:  A  machine  drill  set-up  is  made  next  the 
roof  of  the  sheet,  as  shown  in  Fig.  42,  so  as  to  advance  a  heading  face 
about  7  to  8  ft.  in  height,  leaving  beneath  an  untouched  bench  which 
will  vary  from  10  to  14  ft.  in  thickness  to  the  bottom  of  the  sheet. 
When  the  heading  is  advanced  about  18  ft.  the  bench  beneath  is  drilled 
and  blasted.  The  placing  of  the  holes  in  the  heading  face  is  shown  in 
Fig.  43  (a),  the  firing  being  in  the  order  designated  by  numbers.  All 
the  holes  are  8  ft.  deep,  numbers  1,  2,  and  4  being  drilled  horizontally, 
and  No.  3  bedded  at  the  junction  of  the  roof  and  the  ore.  No.  5  is  given 


MINES  AND    MINERALS. 


FIG.  42. — Blasting  sheet  ground,  S.  W.  Mo. 

a  slightly  downward  pitch.  Placing  3,  4,  and  5  as.  shown  will  usually 
give  a  break  which  leaves  room  for  a  good  set-up.  The  drilling  is  done 
with  a  3  1/4-in.,  type  E  24,  Ingersoll-Rand  drill  set  on  a  7-ft.  column. 
The  long  bench  holes  are  put  in  with  the  drill  mounted  on  a  tripod; 
they  will  average  from  16  to  18  ft.  in  length  and  are  placed  as  shown  in 
Fig.  42.  The  hole  is  then  squibbed  three  or  four  times  with  40  per  cent, 
dynamite  and  finally  the  resulting  chamber  is  filled  with  50  to  100  sticks 
of  the  same  explosive  and  fired  by^a  fuse  and  cap.  Only  the  most 
expert  drillmen  can  put  in  these  long  holes  but  they  have  been  found 
highly  advantageous  because  they  can  follow  one  of  the  compact  rock 
beds.  On  the  contrary,  vertical  holes  across  the  bedding  planes  would 
be  difficult  to  drill  on  account  of  cracks  and  pockets.  For  the  same 


UNDERHAND    STOPING 


97 


reason  they  would  be  so  porous  as  to  render  their  exploding  inefficient. 
It  only  takes  three  of  these  long  horizontal  holes  to  break  off  a  bench 
14  ft.  high  across  a  stope  40  ft.  wide.  Between  two  squibbings  the  hole 
is  cleaned  and  cooled  by  blowing  out  with  compressed  air.  Holes 
are  squibbed  only  at  night  and  those  squibbed  one  night  are  blasted  the 
next  at  the  end  of  the  shift.  A  drill  on  the  bench  can  break  as  fast  as 
four  drills  above  on  the  heading,  even  though  only  one  and  one-half  of 
the  flat  bench  holes  can  be  drilled  in  eight  hours. 

The  illustration  shows  a  hole  after  squibbing,  when  it  is  loaded  with 
from  50  to  150  Ib.  of  40  per  cent,  dynamite  and  fuse-fired.  An  average 
of  35  tons  are  broken  per  50  Ib.  of  powder,  giving  a  cost  per  ton  of  12  cents. 


MINES  AND    MINERALS. 


FIG.  43. — Stope  plan,  S.  W.  Mo. 


The  "dirt"  is  shoveled  into  tubs,  pushed  to  the  shaft,  and  hoisted  in 
the  usual  manner,  using  30x32-in.  tubs  holding  about  950  Ib.  The 
hoisting  is  done  by  friction-geared  hoisters  placed  on  the  floor  of  the 
derrick,  one  above  each  shaft  compartment.  These  are  6x7-in.  engines 
with  24-in.  drums,  of  the  English  Samson  pattern,  manufactured  by  the 
English  Iron  Works,  Kansas  City.  It  takes  45  seconds  for  a  car  to 
make  a  round  trip  with  a  travel  of  about  280  ft.  from  the  derrick  floor 
to  the  shaft  bottom.  The  average  rope  speed  is  about  1,300  ft.  per  min- 
ute, so  that  dumping  a  bucket  requires  about  20  seconds. 

The  total  cost  of  mining  for  the  past  few  months  has  averaged  67 
cents  per  ton  hoisted  including  the  pumping-cost  of  3  cents  per  ton. 

There  is  no  timbering  as  the  roof  is  supported  by  pillars  of  ore  25  ft. 
in  diameter,  set  at  40-ft.  intervals  as  seen  in  the  plan  of  the  mine  work- 


98  MINING    WITHOUT    TIMBEK 

ings,  Fig.  43;  no  definite  pillar  system  is  employed,  as  the  pillars  are  set 
where  the  conditions  of  the  roof  demand.  As  a  rule  the  roof  consists  of 
a  solid  flinty  bedded  rock  averaging  from  18  in;  to  3  ft.  in  thickness. 
It  seldom  causes  trouble  except  where  pitted  with  large  pockets  known 
as  "sand  holes." 

Plans  are  now  afoot  to  make  a  very  radical  change  in  the  method  of 
handling  the  mine  dirt.  An  automatic  traction  shovel  made  by  the 
Thew  Automatic  Shovel  Co.,  of  Dayton,  Ohio,  will  be  installed  with  a 
dipper  of  1/2  cu.  yd.  capacity,  which  will  clean  up  anything  within  a 
circle  of  20  ft.  without  moving  the  shovel.  Power  will  be  supplied  by 
compressed  air.  Cars  will  be  used  instead  of  cans,  and  will  be  hauled 
to  the  shaft  in  trains  by  mules  instead  of  hand  tramming.  Here  at  the 
shaft  it  is  intended  to  put  in  an  underground  hopper  which  will  discharge 
into  3-ton  balanced  skips.  The  hoisting  engine  will  be  moved  to  the 
ground  and  steel  head-frames  substituted  for  the  present  wooden  ones. 
At  present  the  average  daily  mine  output  (two  shifts)  is  about  725  tons 
at  No.  3,  with  a  total  from  the  four  shafts  of  about  2,250  tons.  Hoisting- 
is  now  done  both  day  and  night,  but  it  is  hoped  rthat  the  new  hoisting 
system  will  get  out  enough  dirt  in  one  shift  to  run  the  mill  for  two  shifts. 

At  the  author's  visit,  the  labor  for  the  day's  output  of  1850  buckets 
(879  tons)  was  as  follows: 

Day  shift.  Nightshift. 

32  machine  drillers.  6  machine  drillers. 

20  muckers.  14  Muckers. 

2  shot-firers. 

The  high  average  output  of  26  tons  apiece  for  the  muckers  (which 
included  an  average  tram  of  150  ft.  to  the  shaft  bottom)  was  only  at- 
tained by  allowing  them  to  earn  high  wages  under  individual  contracts. 

Pillar-Robbing. — It.  is  generally  the  practice  in  this  district  to  leave 
the  poor  ore  as  pillars  and  not  attempt  to  recover  it;  but  in  one  mine, 
with  rich  ore,  where  the  roof  was  heavy  and  a  large  amount  of  timber 
had  to  be  used  to  enable  even  half  the  ore  to  be  extracted,  the  following 
method  was  used  'for  robbing  the  pillars.  Beneath  the  ore  was  a  solid, 
compact  limestone  stratum.  The  shaft  was  sunk  a  few  feet  into  this, 
and  a  sub-drift  extended  beneath  the  ore  pillars  with  a  7-  or  8-ft.  roof. 
A  raise  was  put  up  in  the  center  of  each  pillar  and  the  ore  shot  down  into 
the  drift  below,  and  trammed  to  the  shaft.  This  method  gave  a  safe 
place  in  which  to  work  and  at  the  same  time  allowed  nearly  all  the 
ore  to  be  recovered  from  the  pillars. 

GOBBING  AND  SAVING  TIMBERS 

Theoretically  the  gob  should  be  let  in  at  either  end  of  the  section,  at 
the  center,  making  it  possible  to  remove  the  stringers.  The  saving  of 


UNDERHAND    STOPING  99 

stringers  depends  on  how  the  gob  is  let  into  the  stope,  the  weight  of  the 
ground  on  the  section,  and  the  condition  of  the  stringer.  Sometimes  it 
would  cost  more  to  remove  a  timber  than  it  is  worth.  It  such  a  case  no 
attempt  would  be  made  to  save  it. 

In  general,  an  average  of  perhaps  50  per  cent,  of  the  stringers  can  be 
saved  in  a  sulphide  stope,  while  in  an  oxide  stope  with  light  ore  75  to  90 
per  cent,  of  the  stringers  are  saved.  The  square  sets  are  not  gobbed  as 
they  are  used  in  the  mining  of  the  next  section.  Only  the  central  portion 
is  filled.  When  the  next  section  toward  the  main  drift  A  is  mined,  the 
pillar  of  ore  to  be  sliced  is  15x25  feet. 

REQUIREMENTS  FOR  APPLICATION  OP  METHOD 

From  the  method  described,  it  will  be  seen  that  the  requirements  in 
order  to  work  such  a  body  of  ore  are:  (1)  There  must  be  a  solid  back 
which  can  be  easily  supported.  (2)  The  ore  must  contain  little  or  no 
waste,  as  everything  goes  into  the  chutes,  permitting  of  no  selection.  '  (3) 
Lateral  and  vertical  pressure  must  be  small  in  order  to  prevent  the  square 
sets  from  buckling  before  the  stringers  are  put  in;  also  to  allow  the  mining 
of  the  whole  section  before  the  gobbing  is  commenced. 

COST  OP  MINING  REDUCED 

In  regard  to  the  reduction  of  the  cost,  it  is  best  to  compare  this  method 
with  that  which  employs  square  sets  alone.  It  is  evident  that  less  timber 
is  used  with  this  system.  With  the  saving  of  75  per  cent,  of  the  stringers, 
the  working  of  several  sections  alongside  of  each  other  makes  it  necessary 
to  run  only  one  row  of  square  sets  for  each  section  mined.  There  is  a 
saving  of  perhaps  50  per  cent,  of  the  timber  of  that  used  in  square-set 
system.  The  mining  of  the  ore  in  the  square  sets  B  and  Br  would  cost 
approximately  the  same  as  by  the  regular  square-set  system.  The  cost 
of  mining  a  lead  row  of  sets  is  higher  than  mining  corner  sets  in  a  square- 
set  stope,  but  this  increased  cost  is  offset  by  the  fact  that  the  ore  from 
the  lead  row  of  sets  falls  directly  into  the  chutes,  making  shoveling  into  a 
wheelbarrow  and  wheeling  to  a  chute  unnecessary. 

In  mining  the  core,  the  amount  of  powder  used  is  reduced  to  about 
one-half.  The  cost  of  timber  and  timbering  is  also  reduced  to  one-half, 
while  the  cost  of  breaking  ore  is  reduced  to  one-third  of  square-setting. 

SAVING  IN  LABOR 

There  is  a  greater  saving  by  this  system  in  the  handling  of  the  ore  than 
in  the  method  of  timbering.  A  large  percentage  of  the  ore  is  shot  directly 
into  the  chutes  and  requires  little  or  no  handling  except  the  breaking  of 
boulders  which  are  too  large  to  pass  through  the  grizzlies. 


100  MINING    WITHOUT    TIMBER 

In  square  setting  it  is  often  difficult  to  place  the  chutes  so  that  the 
miner  can  shovel  directly  into  them.  With  the  Mitchell  slicing  system 
the  wheelbarrow  is  never  used  and  shoveling  is  reduced  to  a  minimum. 

In  working  out  the  sill  floor,  the  ore  is  handled  by  the  ordinary 
method,  as  here  the  ore  must  be  shoveled  directly  into  the  mine  cars, 
unless  worked  frcm  the  level  below,  which  is  often  done. 

INCREASED  TONNAGE  OBTAINED 

The  amount  of  ground  broken  per  man  per  eight-hour  shift,  when 
using  the  regular  square-set  system,  is  from  5  to  6  tons. 

In  mining  the  pillars  with  the  Mitchell  system  in  sulphide  ore,  12  to 
15  tons  are  broken  per  man  per  shift,  while  in  oxide  ore  in  auger  ground 
25  tons  per  man  per  shift  is  not  unusual.  When  once  the  mining  of  the 
core  commences,  the  work  is  carried  on  rapidly,  a  core  often  being  worked 
out  in  8  or  10  days.  It  has  been  found  convenient  to  mine  these  cores 
when  there  is  any  sudden  demand  for  an  increase  in  the  output  of  a  certain 
kind  of  ore,  which  is  another  valuable  feature  of  the  method. 

CONCLUSION 

The  system  can  be  worked  on  any  section  of  ore  provided  that  it  con- 
tains no  waste,  is  not  too  heavy,  and  is  as  large  as  20x30  ft.  There  is 
flexibility  in  this  method  as  it  may  readily  be  switched  to  square-set 
stoping  in  mining  irregular  portions  of  the  orebody.  It  has  not  been 
found  practical  to  mine  a  section  more  than  50  ft.  thick.  The  system  is 
new,  and  Mr.  Mitchell  is  adding  improvements  which  will  make  it  a  still 
more  valuable  method  of  mining. 

The  method  has  been  a  success,  but  owing  to  its  rigid  requirements, 
its  field  is  quite  small.  It  will  suit  only  a  few  of  the  orebodies  in  Bisbee 
and  therefore  will  not  become  an  important  factor  in  reducing  the  cost 
of  mining  there. 

EXAMPLE  12. — CALUMET  AND  ARIZONA  MINE,  BISBEE,  ARIZ. 
(See  also  Example  23.) 

Underground  Quarrying  of  Panel-cores,  or  the  Mitchell  System  for 
Flat  Lenses  in  Limestone. — Because  of  the  peculiar  conditions  under 
which  most  of  the  orebodies  in  Bisbee  exist,  square-set  stoping  in  panels, 
as  described  in  Example  23,  has  been  the  chief  method  of  extraction. 

In  1908  at  the  Calumet  &  Arizona  mine,  while  working  a  heavy 
sulphide  stope  by  the  square-set  method,  a  large  mass  of  ore  broke  away 
from  the  back,  and  in  order  to  mine  it,  long  timbers  were  thrown  across 
the  top  of  the  ore  to  support  the  back,  after  which  the  ore  was  taken  out. 
From  this  slight  incident  a  combination  of  the  square-set  and  underhand 


UNDEKHAND    STOPING 


101 


stoping  systems  was  worked  out  by  M.  W.  Mitchell,  the  foreman  of  the 
Calumet  &  Arizona  company.  The  system  has  given  excellent  results 
where  the  conditions  have  been  favorable. 

Recently  some  bedded  ore  deposits  have  been  found  in  the  Calumet 
&  Arizona  property  Chalcopyrite,  bornite  and  pyrite  have  replaced 
the  limestone,  the  ore  following  the  original  bedding  of  the  limestone  and 
including  little  waste.  These  bedded  deposits  rarely  exceed  50  or  60  ft. 
in  thickness.  The  limestone  hanging-wall  is  well  defined,  solid  and  easily 
supported.  It  is  in  these  deposits  'that  the  Mitchell  system  has  been 
employed.  The  greatest  success,  however,  has  been  attained  in  the 
mining  of  the  oxide  ores  when  they  contain  little  or  no  waste. 

METHOD  OF  BLOCKING  OUT  THE  ORE 

The  orebody  is  first  thoroughly  prospected  to  ascertain  its  general 
direction,  size,  and  limits,  in  order  to  determine  whether  this  method  is 
suitable.  The  theory  of  this  system  of  stoping  is  to  outline  a  block  of 


:L 


Cut  back 
for  Chute  Jaws 


Leave  Sill  Floor,  to 
be  Mined  from  below 


To  be  Mined  later 


Sill  Floor 


TJie  Engineering  &  Minmy  Jaw-rial 

FIG.  44. — Blocking  out  ore  by  Mitchell  slicing  system. 

ore  by  means  of  regular  square  sets,  allowing  the  included  core  to  rest  on 
its  own  base  and  then  cut  it  out  in  slices  from  the  top  down  after  the  roof 
or  back  has  been  properly  supported.  The  method  followed  is  illustrated 
in  Fig.  44.  Two  lead  rows  B  and  B',  15  ft.  apart,  of  regular  sill-floor-stope 
square  sets,  are  run  from  the  main  drift  A  to  the  end  of  the  section  to 
be  mined.  These  are  connected  by  the  square  sets  C.  Regular  7-ft. 


102 


MINING    WITHOUT    TIMBER 


10-in.  stope  sets  are  carried  up  to  the  limits  of  the  ore  above  the  end  sets 
C  and  above  the  sets  B  and  B'.  These  sets  now  include  on  three  sides  a 
block  of  ore  15x45  ft.  and  as  high  as  the  ore  extends. 

Fig.  45  illustrates  the  method  of  framing  used  for  the  square-set  tim- 
bers. The  posts  and  caps  are  usually  10x10  in.  with  8x10  girts.  In  the 
rows  B  and  Bf  the  ties  or  girts  are  put  in  across  the  drift  with  caps 
running  parallel  to  B  and  B' . 


00      O 

a  -° 
S   2 


I 


With  8x  ]0'Girt 


ap  10  x 


U       Posta  . 
5'c  to  c 


Stope  Set  Timbers 


Post 


m 


4-6- 

Cap 


111 


j 


Girt 

J't.t  Kngineerina  $•  Minin,,  ,/owi-W 

FIG.  45. — Framing  square  sets  at  Calumet  and  Arizona  mine. 


UNDERHAND  STOPING 

The  stoping  system  proper  now  commences  and  is  illustrated  by  Fig. 
46,  which  shows  a  plan  and  two  sections  of  the  stope.  The  drills  are 
mounted  on  columns  or  bars  between  the  caps  or  posts  of  the  square  sets 
and  holes  drilled  from  the  sides.  When  the  ore  is  broken,  stringers  S1 
and  $2  are  Pu^  in  and  Sagamore  or  so-called  segment  sets  S3J  are  put  in 
between  /Sx  and  S2.  In  the  second  slice  or  bench  and  those  following, 
stringers  S4  are  put  in  without  the  segment  sets. 


UNDERHAND    STOPING 


103 


In  mining  the  second  bench,  and  those  below,  the  best  practice  is  to 
mount  the  drill  column  between  the  stringers  and  drill  vertical  holes 
downward.  The  stringers  on  the  top  floor  are  10x10  in.  and  framed  like 
girts  to  fit  the  square  sets.  On  the  second  floor  8xlO-in.  stringers  are 
used,  while  8x8-in.  may  be  used  on  floors  below  provided  that  the  ground 
is  not  too  heavy.  Segment  sets  are  put  in  on  the  top  floor  only  to  support 
the  back.  On  the  remaining  floors  stringers  alone  are  used  with  perhaps 
an  occasional  stull  or  spreader  to  reinforce  them.  The  rows  of  square 
sets  B  and  B'  are  used  as  chutes,  grizzlies  being  put  in  to  prevent  large 
boulders  from  clogging  the  mouth  of  the  chutes  which  are  merely  small 
openings  cut  back  of  every  other  set  on  the  sill  floor  as  shown  in  Fig.  44. 
These  openings  are  cut  just  large  enough  for  a  chute,  when  the  sill-floor 
lead  sets  are  run.  With  a  small  amount  of  barring,  the  cars  are  easily 
loaded  from  these  chutes. 


Waste 


Tke  Engineeri 

Section  C-D 
FIG.  46. — Plan  and  section  of  Mitchell  slicing  system. 

PLACING  OF  TIMBERS 

The  plan  in  Fig.  46  illustrates  some  of  the  details  of  the  method 
employed.  No.  1  shows  the  stringer  in  place.  No.  2  shows  diagonal 
braces  to  hold  the  square  sets  in  position.  No.  3  shows  temporary 
spreads  which  are  sometimes  used  to  reinforce  the  stringers  when  the 
ore  is  blasted.  The  method  of  putting  in  stringers  is  shown  by  No.  4. 
One  end  is  put  in  against  the  posts  and  the  caps  of  the  square  set  in  the 
same  way  an  ordinary  girt  is  put  in.  At  X  one  cap  of  the  square  set  is 
cut  down  2  in.  to  permit  the  2-in.  tenon  of  the  stringer  to  go  into  position. 
When  in  place  a  small  piece  of  plank  is  spiked  to  the  cap  to  hold  the 
stringer.  When  the  section  is  worked  out  and  is  ready  for  gob  vertical 
planking  is  put  on  at  the  end  of  the  section  at  No.  5  in  Fig.  46,  and  the 
inside  of  the  square  sets  is  lagged,  as  shown  by  No.  6.  When  this  has 


104 


MINING    WITHOUT    TIMBER 


been  done  with  the  ore  worked  out  to  the  level  or  to  the  bottom  of  the 
orebody  the  stope  is  ready  for  gob. 

EXAMPLE  13. — SECTION — 21  MINE,  ISHPEMING,  MARQUETTE  RANGE,  MICH 

(See  also  Example  35.) 

Underground  Milling  in  Sub-vertical  Vein  with  Back  Caving. — The 
Section  21  mine  (Oliver  Co.)  is  3  miles  south  of  Ishpeming  adjoining 
the  Whitby  open  pit,  which  was  exhausted  some  years  ago  by  surface 
milling,  and  used  an  inclined  skip  for  hoisting  the  ore.  The  west  end  of 
the  present  mine  is  a  large  trough  of  continuous  non- Bessemer  soft  ore, 
which  is  worked  by  the  room-caving  system  of  Example  46.  On  the  east 


Plan 


Long.  Sec. 
FIG.  47. — S toping  at  Section— 21  mine. 

end  the  ore  is  mediumly  hard  and  has  a  thickness  of  15  ft.  to  50  ft.,  lying 
in  a  highly  inclined  lense  on  a  diorite  footwall  with  a  jaspilite  hangwall 
and  is  worked  by  underground  milling. 

In  Fig.  47  the  drift  d  of  the  new  level  should  be  completed  as  soon 
as  the  stope  above  level  D  (60  ft.  higher)  has  reached  the  contour 
mnpqr.  Raises  b  and  e  (50  ft.  apart)  are  then  put  up  from  d  to  D.  After 


UNDERHAND    STOPING  105 

leaving  a  6-ft.  pillar  K  under  level  D  breast  slopes,  as  s-sl  and  s2-s3  are 
begun  under  the  pillar  K  in  each  raise  in  order  to  start  the  ordinary 
annular  underhand  benches  v,  vl,  etc.,  of  milling,  these  being  cut  down 
till  the  limiting  contour  for  self-clearing  (a  b  c  e  /)  is  reached. 

The  robbing  of  the  pillars  m  p  r  still  above  level  D  can  then  begin  by 
putting  upraises  n1  n  and  q1  q  into  the  highest  portions  and  milling 
down  around  these  raises  into  the  chutes  6  and  e,  which  spout  into  a 
tram  car  c  in  d.  As  such  ore  pillars,  when  cut  by  a  dike,  are  liable  to 
slip  along  the  plane  of  contact  it  is  necessary  to  begin  their  extraction  at 
the  end  of  the  lense  at  m,  in  order  to  minimize  the  danger  to  the  miners 
of  a  collapse  of  the  hangwall. 

Finally,  only  enough  of  a  pillar  remains  above  level  D  to  sustain  the 
filling,  and  this  pillar  is  drilled,  as  is  also  the  pillar  K.  The  next  step  is 
to  blast  both  these  drilled  pillars  (continuously  by  fuse-firing)  until  all 
the  ore  remaining  above  the  contour  a  b  c  e/has  sunk  into  chutes  b  and  e. 
It  is  true  that  the  filling  also  descends,  but  most  of  the  ore  can  be  drawn 
from  the  chutes  before  the  filling  appears. 

In  one  case  an  ore  lense  extending  150  ft.  above  the  level  was  worked 
as  one  mill  by  putting  a  raise  to  the  top  and  starting  the  mill  at  that 
point.  As  the  dip  of  the  footwall  had  a  pitch  of  45  deg.  it  was  easy  to 
stand  on  it  and  inspect  the  hangwall  to  guard  against  accident  to  the 
miners  who  were  cutting  down  the  benches. 

This  system  can  only  be  safely  worked  where  the  ore  and  hangwall 
are  strong  enough  to  sustain  themselves  over  the  width  of  the  vein  and 
for  the  height  and  length  required  for  economical  milling.  A  weak  back, 
however,  could  be  sustained  over  the  underhand  stope  by  a  V  arch  of 
heavy  stulls  or  "saddle  back"  which  was  formerly  used  with  success  at 
the  Fayal  mine  on  the  Mesabi  to  cover  rooms  23  ft.  wide  and  60  ft.  high, 
whose  sides  were  untimbered.  The  footwall  must  also  be  steep  enough 
to  clear  itself  by  gravity.  Several  levels  can  often  be  worked  simul- 
taneously by  postponing  the  robbing  of  the  pillars  until  later.  The 
ventilation  is  good  and  little  or  no  timbering  or  shoveling  is  required. 
The  ore  is  not  broken  as  ordinarily  in  caving  systems,  chiefly  by  hang- 
wall  pressure,  but  the  next  cheapest  method,  i.  e.,  underhand  stoping, 
is  employed,  which  requires  few  expensive  development  openings. 

In  suitable  ground  the  chief  objection  is  in  the  loss  of  ore  through 
contamination  by  the  filling;  but  this  does  not  preclude  its  use  in  the 
mining  of  many  iron  ores.  It  is,  also,  only  adapted  to  orebodies  that  are 
homogeneous,  as  no  sorting  can  be  conveniently  done  underground. 


CHAPTER  IX 
OVERHAND  STOPING  WITH  SHRINKAGE.     NO  FILLING 

EXAMPLE  14.     WOLVERINE  COPPER  MINE,  HOUGHTON  COUNTY,  MICH 

(See  also  Example  19.) 

Sloping  Amygdaloid  Beds  with  Strong  Walls  (No  Chutes). — For  the 
amygdaloids  and  conglomerates  of  this  district,  the  mining  problem 
is  to  excavate  practically  the  whole  contents  of  beds,  from  3  to  30  ft. 
thick,  of  indefinite  depth,  and  of  a  length  along  the  strike,  depending  on 
the  mineralization,  but  seldom  less  than  several  hundred  yards. 
Usually  the  beds  have  a  greater  dip  than  the  angle  of  repose  between 
broken  rock  and  footwall. 


FIG.  48. — Shaft  station  in  Wolverine  mine. 

In  developing  the  Wolverine  (Fig.  48)  amygdaloid  with  its  strong 
hangwall  the  drifts  are  at  100-ft.  intervals,  and  are  carried  20  ft.  high 
across  the  vein,  for  the  length  of  the  payshoot,  which,  including  barren 
spots,  is  3000  ft.  An  overhand  stope  is  then  started  above  a  drift  (see 
Fig.  49)  and  extends  up  to  the  10-ft.  longitudinal  rib,  under  the  level 
above.  In  the  excavation,  the  only  hanging  wall  support  is  a  pillar  p 
(15  ft.  diameter),  for  each  75-ft.  room;  which  is  formed  by  cutting  around 
until  only  a  .10-ft.  neck  of  ore  is  left  which  can  be  pierced  by  a  single 
round  of  drill  holes.  In  long  ore  shoots  it  is  best  to  also  leave  a  15-ft. 
panel  pillar,  along  the  dip  from  level  to  level,  every  three  or  four  rooms. 

106 


OVERHAND    STOPING    WITH    SHRINKAGE 


107 


A  room  is  let  on  contract  to  four  men  (two  on  each  shift  with  one 
air  drill)  with  quarterly  settlements,  the  monthly  advance  to  each  man 
being  $65.  To  simplify  the  calculation  of  contract-excavation,  the 
vein  is  assumed  to  have  an  average  width  of  2  fathoms,  so  that  only 
the  distance  stoped  along  the  drift  and  up  the  footwall  need  be 
measured.  For  drifting,  a  sectional  area  of  2  sq.  fathoms  is  deducted 
from  the  stope,  and  this  is  paid  for  at  the  rate  of  $5  to  $5.50 
per  lineal  foot,  while  the  stoping  itself  is  let  at  $7  to  $9  per  cubic 
fathom  (216  cu.  ft).  At  these  prices,  everything  is  furnished  by  the 
company  except  explosives  (the  powder  being  charged  $17  for  a 
50-lb.  box).  One-third  the  wages  ($30  a  month)  of  the  nipper  boy 
and  the  wear  of  steel  is  also  paid  by  the  contractors.  For  wear  of  steel, 
each  man  is  charged  $1  a  month,  and  in  the  quarterly  settlement,  any 
ost  drills  are  put  in  at  the  rate  of  25  cents  per  Ib. 


FIG.  49. — Stoping  at  Wolverine  mine. 

Shaft-sinking  is  also  let  on  similar  contracts,  the  price  being  around 
$16  per  lineal  foot  for  a  section  8x17  ft.  in  the  clear.  To  prevent  sub- 
letting, all  contractors  are  paid  individually. 

The  muckers  are  paid  $2.30  a  shift,  and  work  in  pairs,  each  pair 
having  a  stunt  of  loading  and  tramming,  from  stope  to  shaft-station, 
40  2-ton  cars  per  shift.  The  footwall  slope  of  40  deg.  is  sufficient  to 
cause  the  coarser  broken  ore  to  roll  into  the  drift,  whence  it  is  shoveled 
into  cars,  as  no  chutes  are  put  in.  The  waste  from  dead  work  is  dumped 
into  old  stopes,  though  recently  a  little  has  been  used  for  dry-walling 
to  support  some  weak  hanging  wall.  Only  a  small  percentage  of  broken 
lode  need  be  rejected  in  the  rock-house  above,  as  too  poor  for  the  mill. 

The  stopes  are  cut  out  in  the  usual  horizontal  benches  B  (Fig.  49) 
and  water  holes  are  drilled  wherever  possible.  Six  or  seven  of  8-  to  10-ft. 
holes  can  be  bored  per  shift  with  the  No.  3  Rand  drill  in  use.  This 


108  MINING    WITHOUT    TIMBER 

permits  an  excavation  of  35  to  50  cubic  fathoms  per  month  per  machine, 
or  about  1000  tons  of  broken  ore.  Twelve  pounds  of  40  per  cent, 
dynamite  will  break  a  cubic  fathom  of  ore.  A  raise  R  is  put  through  the 
longitudinal  rib  at  each  room  for  ventilation. 

At  my  visit,  37  drills  were  used  on  the  day,  26  on  the  night  shift. 
The  drills  are  sharpened  by  a  Ward  Bros,  machine,  with  a  coke-heater 
and  a  forge  blown  by  a  special  fan  made  by  the  Garden  City  Fan  Co. 
of  Chicago.  The  bits  have  +  points  up  to  4-ft.  length,  and  beyond  that, 
chisel-points.  The  drills  are  generally  run  from  tripods,  set  on  a  scaffold, 
or  partly  on  a  stull  in  the  wider  stopes. 

For  executives  there  are  a  captain  and  assistant  (both  on  day  shift) 
and  a  night  shift  boss,  whose  chief  duties  are  to  see  that  only  good  ore 
is  broken  down.  Unprofitable  portions  of  the  vein  are  left  as  extra 
pilars,  but  the  stope  contractors  are  allowed  something  extra  at  the 
settlement  for  their  consequent  loss  in  volume,  as  is  also  the  case  if  the 
stope  has  exceeded  the  assumed  average  thickness  of  2  fathoms.  For 
the  muckers,  there  is  a  boss  on  each  shift  at  every  working  shaft. 

This  mining  system  as  described  is  suited  to.  an  ore-bed  compara- 
tively free  from  waste  and  having  a  hangwall  strong  enough  to  stand 
alone  over  wide  areas  and  a  footwall  sufficiently  steep  to  be  self-cleaning. 

EXAMPLE  15. — HOMESTAKE  MINE,  BLACK  HILLS,  SOUTH  DAKOTA. 

Sub-vertical  Wide  Vein  With  Strong  Walls,  No  Chutes. — Many  of 
the  early  miners  at  the  Homestake  were  from  Virginia  City,  Nev.,  and 
as  in  a  great  many  other  camps,  the  early  mining  is  a  record  of  Comstock 
methods — a  desire  to  square-set  everything.  Even  to-day  there  are 
timbered  stopes  still  unfilled,  and  which  will  stand  open,  no  doubt,  long 
after  the  square-set  timbers  rot  and  fall  apart.  Later,  when  the  system 
of  "open-stope"  for  filled-stope  mining  was  adopted,  still  clinging  to  the 
old  idea  of  putting  in  the  timber,  the  entire  sill  floor  was  square-set, 
supposedly  for  the  purpose  of  keeping  the  haulage  gangways  from  swing- 
ing. With  this  method  no  lagging  was  used  over  the  timbers,  except  to 
protect  the  gangways,  and  it  is  said  that  the  amount  of  timber  that  was 
crushed  and  broken  in  filling  the  sill  floor  was  enormous.  When  the 
stope  was  drawn,  this  timber  caused  endless  trouble. 

The  uselessness  of  timbering  anything  other  than  the  haulage  ways  on 
the  sill  floor  was,  of  course,  soon  evident,  so  the  next  change  in  method 
was  to  break  out  the  sill  floor,  then  shovel  through  the  necessary  gang- 
ways, timbering  and  lagging  them,  and  packing  rock  around  them  as  a 
protection;  the  stope  was  then  carried  up.  This  method  is  still  used  in 
many  places,  but  the  last  stage  in  the  development  of  the  stoping  practice 
has  been  to  do  away  with  the  use  of  even  this  timber  wherever  possible. 

The  main  ledge  is  so  wide  throughout  most  of  its  extent  that  the  stopes 


OVERHAND    STOPING    WITH    SHRINKAGE 


109 


must  be  carried  across  the  orebody  instead  of  along  its  strike.  In  other 
words,  the  hanging-and-foot  walls  are  the  ends  of  the  stopes  instead  of 
their  side  walls.  The  width  to  which  stopes  are  best  worked  parallel  to 
the  strike  varies  with  the  nature  of  the  wall  rocks,  but  it  may  be  stated 
that  as  a  general  thing  when  the  ledge  attains  a  width  of  more  than  80  ft., 
it  has  been  considered  best  to  lay  out  the  stope  across  the  orebody.  In 
many  places  the  orebody  is  over  400  ft.  wide.  The  No.  1  North  stope 
on  the  700-ft.  level  was  60x520  ft.  on  the  sill  floor.  This  stope  was 
worked  with  square  sets.  In  working  the  level  above  the  900,  stopes 
were  carried  60  ft.  wide  from  foot-  to  hanging-wall,  and  pillars  of  60-ft. 
width  were  left  between  stopes.  More  recently,  however,  60-ft.  stopes 
and  42-ft.  pillars  have  been  adopted;  the  1500  level  from  the  Ellison  shaft 
is  being  laid  out  on  this  plan.  To  a  depth  of  1100  ft.,  levels  were  carried 
at  100-ft.  intervals;  below  that,  they  are  150  ft.  apart. 

PRESENT  STOPING  SCHEME 

The  usual  method  of  approaching  these  cross  stopes  through  timber- 
less  crosscuts  is  shown  in  Figs.  50  and  51.  The  ore  is  drawn  into  these 
crosscuts,  shoveled  into  cars  and  trammed  out  in  timbered  drifts. 


FIG.  50. — Cross-section  of  stope,  Homestake  mine. 

The  orebody  is  first  developed  by  a  drift,  as  shown.  Laterals,  or 
crosscuts,  are  then  turned  off  at  102-ft.  centers  and  run  through  to  the 
walls.  Simultaneously,  the  stope  sills  being  cut  out  by  driving  across 
between  the  pillar  crosscuts  and  breaking  out  to  the  full  60-ft.  width. 
The  crosscuts  pass  through  the  center  of  the  pillars,  and  at  30-  to  35-ft. 
centers  connections  to  serve  as  draw  holes  are  broken  through  to  the  stope. 
These  crosscuts  are  connected  with  the  footwall  drifts,  serving  as  main 
haulage  ways.  The  stopes  are  then  worked  up,  just  enough  ore  being 
drawn  so  as  to  keep  the  drillers  within  reach  of  the  back.  One  or  more 
manways,  depending  upon  the  conditions,  is  carried  up  with  each  stope. 


110 


MINING    WITHOUT    TIMBER 


By  maintaining  the  footwall  drifts  and  tapping  stopes  through  crosscuts 
in  the  pillars,  timbering  is  practically  eliminated  in  the  first  mining  stage. 
The  stopes  are  usually  carried  up  to  within  20  ft.  or  so  of  the  level 
above;  the  back,  or  crown,  being  removed  after  the  stope  has  been  com- 
pletely emptied  of  ore  and  filled  with  waste.  The  crowns  are  taken  up 
in  small  sections  of  24  or  30  ft.,  using  square-set  timbers.  On  the  300 
level,  No.  1  Pierce  stope,  the  crown  was  being  taken  out  at  the  time  of 
my  visit.  This  stope  is  about  100  ft.  wide  and  200  ft.  long,  and  the  crown 
was  probably  30  ft.  thick.  A  hole  was  broken  through  on  the  footwall, 
where  the  ore  is  generally  of  better  grade  and  the  rock  benched  back 


Fermanent  Foot  Wall  Drjft 


-Slate- 


FIG.  51. — Plan  of  stope,  Homestake  mine. 

toward  the  hanging.  Finally  about  three  sets  of  timber  were  put  in  next 
to  the  footwall,  and  under  the  remainder  of  the  crown,  which  was  then 
carefully  worked  out  to  the  hangingwall.  In  breaking  out  the  crown,  the 
ore  is  left  across  both  ends  of  the  stope  so  as  to  form  a  supporting  arch, 
until  the  timbers  are  under  the  hangingwall  portion.  This  work  is 
dangerous  and  requires  careful  watching. 

It  is  planned  to  take  out  the  pillars,  even  in  the  lower  workings,  by 
working  them  in  small  sections  of  square-set  timbered  stopes.  This  will 
be  the  second  stage  of  mining  at  the  Homestake.  After  the  ore  is  drawn 
from  the  primary  stopes  the  sides  against  the  pillars  are  laced  up  with 
lagging  set  vertically,  to  which  slabs  laid  horizontally  are  nailed.  A 
section  of  lacing  is  put  up,  then  waste  run  in  from  above  until  this  section 


OVERHAND    STOPING    WITH    SHRINKAGE  111 

is  filled;  another  section  of  lacing  is  placed,  more  waste  run  in,  etc. 
Doubtless,  by  the  time  the  pillars  are  removed  the  slabs  will  in  many 
cases  be  rotted,  but  they  serve  to  catch  up  the  waste  as  square-sets  are 
put  in  the  pillar  or  secondary  stopes.  In  all  cases  the  crown  over  the 
original  stopes  must  be  taken  out  before  the  pillars  are  worked,  or  else 
this  ore  would  probably  be  lost  by  caving,  there  being  no  support  on  the 
sides  of  the  stope. 

BREAKING  THE  ORE 

Up  to  date,  only  large  piston  drills  have  been  used  for  breaking  ore  in 
the  Homestake  mine.  (Trials  are  now  being  run  with  several  makes  of 
stoping,  air-hammer  drills.)  By  putting  in  long  holes  and  picking  favor- 
able places,  huge  masses  of  rock  ore  slabbed  down.  It  is  this  tendency 
of  the  ore  to  break  large  that  accounts  for  the  great  amount  of  shoveling 
necessary.  The  ore  will  not  run  through  chutes,  and  at  each  gate  block- 
holers  with  "Jap"  plugger  drills  are  kept  busy  drilling  and  breaking  the 
ore  so  that  it  can  be  handled  into  the  cars.  About  3  Ib.  of  No.  2  dyna- 
mite is  consumed  per  ton  of  ore  placed  in  the  mine-car.  Labor  at  the 
Homestake  costs  $3  per  day  for  trammers  and  shovelers,  and  $3.50  for 
machine  men.  The  labor  union  is  not  recognized. 

Conditions  at  the  Homestake  are  ideal  for  the  operation  of  shrinkage 
stopes,  the  ore  being  tough  enough  to  present  a  back  under  which  the 
men  may  work  with  safety  and  the  walls  being  good  and  tight.  Caving 
is  the  only  other  mining  method  that  might  seem  applicable  for  working 
such  an  immense  low-grade  deposit  at  a  profit.  This,  however,  is  not 
feasible  as  the  ore  is  too  tough  and  hard.  On  the  lower  levels  the  ore  is 
ahornblendic  schist  containing  much  fine  disseminated  iron  pyrite  and 
this  so  increases  its  specific  gravity  that  only  10  cu.  ft.  weigh  a  ton. 
In  places  near  the  surface  immense  portions  of  the  orebody  have  been 
broken  away,  and  after  years,  of  crushing  and  packing,  are  not  yet  suffi- 
ciently broken  up  so  that  the  ore  will  run  in  chutes.  In  these  places  a 
small  square-set  stope  is  run  up  a  few  sets,  a  grizzly  put  in  at  the  top  and 
rock-blasted  down,  being  run  for  waste  or  ore,  according  to  its  character. 

DISCUSSION  OF  SYSTEM 

The  objection  to  this  method  is  the  excessive  amount  of  shoveling 
necessitated.  Every  bit  of  broken  ore  is  mucked  into  cars  by  hand,  and 
this  alone  means  a  cost  of  close  to  20  cents  per  ton  of  rock  handled.  The 
cost  of  labor  amounts  to  almost  three-fourths  of  the  total  mine-operating 
cost.  Many  efforts  have  been  made  to  overcome  this  excessive  labor 
consumption  for  shoveling,  but  as  yet  no  satisfactory  solution  has  been 
reached.  However,  just  now  over  one-half  of  the  ore  is  recovered  without 
timbering,  whereas  formerly  everything  was  worked  with  square  sets. 


112 


MINING    WITHOUT    TIMBER 


EXAMPLE  16. — GRATZ  LEAD  MINE,  OWEN  COUNTY,  KENTUCKY 

Sub-vertical  Narrow  Vein  with  Strong  Walls,  Rill  Chutes. — The 
geological  formation  is  the  Cincinnati  of  the  lower  Silurian  period  and 
around  the  mine  comprises  a  limestone,  horizontal  and  thinly  bedded. 

The  vein  dips  nearly  vertically  and  appears  to  be  a  fissure  crack  due 
to  folding,  and  faulting,  if  it  took  place,  must  have  been  along  the 
strike  rather  than  along  the  dip.  The  filling  is  barite,  calcite,  and  galena: 
the  first  occurs  in  typical  white  or  brownish  orthorhombic  prisms,  the 
second  in  translucent  white  or  yellow  tablets,  while  the  galena  is  in  cubes, 
either  in  thin  bands  parallel  to  the  walls  or  in  isolated  crystals. 

The  vein  walls  exhibit  no  slickensides  or  other  signs  of  movement,  but 
the  filling  is  frozen  to  them  as  when  first  deposited. 


O  ^y  O    "W  O    N^    O     ^    O     Xp' 
^\/'  oSxx'     O     x>x    O  N^''     O     s--'    O    \ 


FIG.  52. — Stoping  at  Gratz  mine. 

On  the  first  level,  where  the  stoping  is  almost  continuous  for  500  ft., 
the  vein  thickness  varies  from  6  in.  to  6  ft.,  with  an  average  of  above 
15  in.  A  3-ft.  width  of  stope,  however,  must  be  excavated  for  machine 
drilling  and  all  the  broken  rock  is  run  through  the  mill. 

The  mine  is  opened  by  three  levels,  placed  100  ft.  apart  on  the 
5x8-ft.  vertical  shaft,  that  is  only  timbered  through  the  surface  soil.  For 
drilling  the  stopes  there  are  three  2  3/4-in.  Rand  drills  and  the  stoping 
system  is  shown  in  section  along  the  vein  in  Fig.  52. 

Wooden  chutes  are  placed  at  a,  6,  c,  and  d,  from  40  to  50  ft.  apart, 
and  the  stope  bottom  carried  up  hopper-shape,  as  shown,  so  as  to  leave 
pillars  or  "rills"  like  abf  to  protect  the  drift  and  to  avoid  the  use  of 
timber.  These  chute  pillars  can  be  finally  recovered  by  underhand 
stoping.  In  order  to  save  set-ups,  the  stope  back  is  attacked  by  the 
sawtooth  system.  In  stope  1-6,  a  3-ft.  drill  bar  would  be  set  up  for  a 


OVERHAND    STOPING    WITH   SHRINKAGE  113 

certain  round  only  at  points  1,  2,  and  6,  and  the  drill  at  2  would  bore  two 
holes  above  and  two  holes  below  the  bar  in  direction  2-7,  and  then  four 
more  in  direction  2-8,  or  eight  in  all.  Holes  from  points  3,  4,  etc.,  would 
be  put  in  similarly. 

For  the  next  upward  round  the  bar  would  be  set  up  at  points  7,  8,  9, 
10,  and  11,  with  eight  holes  to  be  bored  from  each.  In  this  way  all  the 
holes  are  self-cleaning  uppers  and  a  6-ft.  depth  can  be  reached  without 
sticking.  With  the  flat  and  down  holes  of  usual  overhand  benches,  the 
calcite  cubes,  chipped  out,  would  tend  to  wedge  the  bit  and  cause  delay. 

Enough  broken  ore  is  left  in  the  stope  to  support  the  men  at  the  back, 
for  whose  ingress  plank  mainways  a-2,  6-5,  etc.,  are  carried  up,  with 
entrances  next  the  chute  gates.  For  stoping  the  30  to  40  per  cent, 
dynamite,  used  in  development,  has  been  replaced  by  15  per  cent.,  as 
the  latter  is  slower  and  makes  less  galena  fines  for  the  mill.  The  mine 
only  runs  a  10-hour  day  shift  and  the  drillmen  only  bore,  all  loading  and 
firing  being  done  by  an  extra  gang  at  night. 

This  system  is  well  adapted  to  narrow,  steep  veins  with  hard  walls 
where  no  sorting  of  waste  need  be  done  in  the  stopes.  The  rills  at  the 
chutes  save  timber  and  the  breaking  of  the  back  in  sawtooth  profile  is 
designed  for  boring  the  most  holes  with  the  fewest  set-ups  of  the  air-drill 
bar  where  upward-pointing  holes  are  the  most  advantageous. 

EXAMPLE  17. — ALASKA  TREAD  WELL  MINE,  DOUGLAS  ISLAND,  ALASKA 

(See  also  Example  9.) 

Sub-vertical  Wide  Vein  with  Strong  Watts.  Rill  Chutes. — For  the 
underground  work  the  main  crosscut  C  (Fig.  38)  and  the  drifts  D  are 
arranged  as  for  the  opencut  system.  In  addition,  at  the  ends  of  crosscuts 
C  and  200-ft.  to  500-ft.  intervals  along  the  deposit,  the  different  levels 
are  connected  by  raises  for  manway  and  ventilation  purposes.  The 
drifts  and  crosscuts  are  10x7  ft.  and  the  raises  are  6x8  ft.  in  the  clear,  no 
timber  being  needed  for  either  in  this  hard  formation,  which  is  so  free 
from  seams  and  so  difficult  to  break  that  cut-holes  for  the  drives  must  be 
pulled  with  70  per  cent,  dynamite. 

Stoping  System. — As  the  value  of  the  ore  does  not  permit  of  timbering 
or  filling,  the  present  successful  system  dispenses  with  both.  The  object 
of  the  intermediate  drift,  E,  is  to  open  communication  with  the  ore- 
chutes  and  to  furnish  a  large  facial  area  for  the  machine-drills  to  work 
upon,  in  cutting  out  or  under-cutting  the  ground-floor  for  the  stopes. 
When  the  intermediate  has  advanced  about  50  ft.  the  work  of  cutting 
out  the  stope  is  started.  This  consists  of  mining  out  a  chamber  7  ft. 
high,  from  150  to  300  ft.  long,  and  with  a  width  varying  with  the  width  of 
the  orebody.  In  the  past  it  has  been  customary  to  cut  the  stopes  with 
a  level  floor,  but  experience  has  shown  that  it  is  more  economical  to  cut 
the  floor  so  that  it  slopes  up  from  drifts  D  at  an  angle  of  about  30  deg. 


114  MINING    WITHOUT   TIMBER 

This  does  away  with  a  large  amount  of  shoveling,  and  the  sawtooth 
stope-floor  of  ore  thus  left  at  H  is  ultimately  obtained  through  the 
stopes  from  the  next  lower  level. 

When  the  stope-floor  has  been,  cut  out,  the  work  of  stoping  upon 
the  ore  is  immediately  begun.  The  roof  of  the  stope  is  arched  across- 
from  wall  to  wall  of  the  lode,  thus  serving  the  double  purpose  of  sup- 
porting the  back  and  offering  a  better  surface  for  the  attack  of  the 
machine-drills.  The  ore  is  shot  down  in  large,  thin  slabs,  so  that  the 
shock  of  falling,  combined  with  that  of  the  blasting,  breaks  it  up  as 
much  as  possible.  The  pieces  of  rock  too  large  to  pass  through  the  ore- 
chutes  are  broken  by  hand  and  "  bull-dozed "  with  powder  to  the  re- 
quired size.  When  starting  from  the  floor,  the  machine-drills  begin  in 
the  drift  midway  between  the  lode  walls  and  cut  out  a  trench  along  the 
center  of  the  back  to  form  the  apex  of  the  arch,  its  height  varying  with 
the  character  of  the  rock.  Two  sizes  of  machine-drills  are  used:  the 
3  1/4-in.  and  3  5/8-in.  Ingersoll-Sergeant,  and  the  holes  are  drilled  to  an 
average  depth  of  8  ft.  A  machine-stoping  will  drill  an  average  of  28 .69 
ft.  per  shift  of  10  hours  and  break  34 . 96  tons  of  ore  with  the  consump- 
tion of  12.53  Ib.  of  No.  2  dynamite.  The  cost  o'f  blasting  up  the  rock 
after  it  has  been  blasted  down  is  a  large  item  in  the  expense  of  stoping. 
One  rock-breaker  is  usually  required  to  each  machine,  and  it  takes 
0.85  Ib.  of  powder  in  "bulldozing"  for  each  ton  of  rock  broken. 

As  no  timber  is  used,  it  is  compulsory  that  a  sufficient  quantity  of 
broken  ore  be  left  in  the  stopes  to  form  a  solid  working-floor  for  the 
miners.  It  has  been  found  that  one-third  of  the  broken  ore  can  be  drawn 
off  while  the  stope  is  being  worked,  and  the  surface  of  the  broken  ore 
kept  within  working  distance  of  the  back.  In  other  words,  by  the  above 
methods,  two-thirds  of  the  ore  broken  must  be  left  in  the  stope,  and 
cannot  be  drawn  off  until  the  stope  is  worked  up  to  the  next  higher  level 
and  finished.  In  the  Treadwell  mine  the  slate-horse  forms  a  natural 
division  between  the  stopes  of  the  north  and  the  south  orebodies.  The 
walls  of  the  orebody  are  supported  by  vertical  pillars,  or  ribs,  15  ft. 
thick,  and  from  200  to  300  ft.  apart.  For  means  of  communication  and 
ventilation,  man-way  raises  are  put  in  these  pillars  and  connected  with 
the  levels.  At  intervals  of  25  ft.,  short  drifts  are  run  in  opposite  direc- 
tions from  the -man-way  raise;  so  that,  as  the  working-floor  of  the  stope 
advances,  each  of  them  is  used  successively  when  the  workings  connect 
with  the  main  raise,  and  in  turn  abandoned  and  closed  up  as  connec- 
tion is  made  with  the  next  higher  one.  The  levels  are  protected  by  hori- 
zontal pillars  from  20  to  30  ft.  thick.  Heretofore,  these  pillars  have  been 
left  in  place;  yet,  even  with  this  saving,  fully  20  per  cent,  of  the  ore  must 
remain  in  the  mine  in  the  shape  of  pillars  and  ribs  to  support  the  ground 
and  to  prevent  caving. 

Samples  and  Maps. — Close  attention  is  paid  to  sampling  and  record- 


OVERHAND    STOPING    WITH   SHRINKAGE  115 

ing  the  assay  value  of  the  ore.  As  a  drift,  raise,  cross-cut,  or  other 
development-work  is  in  progress,  a  sample  is  taken  after  each  round  has 
been  blasted.  These  samples  are  taken  either  by  the  shift-boss  or  the 
foreman,  and  their  description  and  location  are  recorded  on  a  special 
tag,  enclosed  with  the  sample  in  the  sack. 

At  intervals  of  15  ft.,  and  closer  if  there  is  any  doubt  as  to  the  value  of 
the  ore,  lines  of  samples — each  sample  being  10  ft.  long  and  varying  with 
the  nature  of  the  ore — are  taken  across  the  back  of  the  stopes  at  right- 
angles  to  the  strike.  These  samples  are  taken  by  cutting  trenches, 
usually  10  ft.  long,  4  in.  wide  across  the  strike  of  the  ore,  and  5  ft.  apart, 
for  the  entire  length  of  the  new  work.  A  hand-sample  is  taken  from  each 
car  at  the  ore-bins,  and  again  at  the  crushers  a  grab-sample  is  taken  by 
means  of  large  dippers,  before  the  ore  goes  to  the  mill.  When  the  mine- 
sample  reaches  the  assay  office,  it  weighs  from  50  to  150  Ib. 

A  complete  set  of  maps  is  kept,  showing  in  detail  the  underground 
and  surface  workings  of  the  mines,  also  the  value  and  position  of  each 
sample  taken  and  the  quantity  of  broken  ore  and  reserves. 

Labor. — On  account  of  the  system  adopted  for  working  the  mines,  due 
to  the  character  of  the  walls  and  vein-material,  it  is  necessary  to  employ 
only  skilled  labor  in  the  shafts,  drift,  raises,  etc. 

About  60  per  cent,  of  the  machine-men  and  helpers  on  the  island  came 
as  laborers  and  have  learned  their  trade  here.  They  are  preferred  by 
the  foremen  and  seem  on  the  average  to  break  more  rock  than  miners 
who  have  learnt  the  trade  elsewhere.  Machine-men  get  from  $2.50  to 
$3 . 00  and  muckers  $2 . 00  per  10-hour  day  with  board  and  lodging. 

This  mining  system  is  suitable  for  wide,  steep  veins  with  strong  back 
and  walls  where  no  waste  need  be  sorted  out  in  the  stopes. 

EXAMPLE  18. — VETA  GRANDE  MINE,  CANANEA,  SONORA,  MEXICO 
(See  also  Examples  6,  34  and  45.) 

Sub-vertical  Lenses  in  Porphyry  (Panel  Cores  with  Pyramidal  Rill 
Chutes. — This  system  of  stoping  combines  square  setting  and  overhand 
stoping  on  ore.  On  the  main  level  the  ore  is  first  blocked  out  with  a 
series  of  drifts  at  right  angles  to  each  other,  one  way  the  drifts  being  40 
ft.  apart,  and  the  other  way  50  ft.  apart,  center  to  center.  The  general 
appearance  resembles  a  checkerboard.  All  the  drifts  are  timbered 
with  regular  sill-floor  stope  square  sets.  Chutes  are  put  in  every  other 
set.  On  the  next  floor  above  the  drift  regular  stope  square  sets  are  put 
in  and  the  square-set  chutes  are  carried  up  one  floor.  On  the  third 
fioor,  that  is  16  ft.  above  the  rail,  the  square  sets  are  put  in  above  the 
row  in  the  drifts  only  and  the  included  rectangle  is  mined  out  on  this 
floor.  From  here  up  this  continues  with  the  square  sets  and  the  chutes 
carried  up  slightly  in  advance  of  the  central  portion  of  the  rectangle. 

Enough  ore  is  drawn  off  through  the  chutes  to  give  the  miners  sufficient 


116 


MINING    WITHOUT    TIMBER 


head-room  to  work  on  the  ore.  In  this  way  these  different  rectangles 
outlined  by  square  sets  are  carried  up  to  the  limits  of  the  orebody. 
There  are  several  kinds  of  chutes  that  can  be  used,  and  it  is  not  necessary 
to  carry  up  a  regular  square-set  chute.  A  simple  beveled  plank  chute 
is  just  as  good  and  uses  less  timber.  In  mining  one  of  these  rectangles 
the  back  is  filled  with  holes  and  all  fired  together.  If  there  is  a  horse  of 
waste  in  the  ore  it  can  be  easily  removed  and  dropped  into  the  chutes 
and  trammed  away.  A  large  amount  of  waste  is  left  in  pillars.  The 
rows  of  square  sets  are  lagged  on  the  outside,  holding  the  ore  in  the  center 
of  the  rectangle  until  the  drawing  commences. 

DRAWING  THE  CORE 

After  the  whole  body  has  been  worked  out  in  this  way,  the 
ore  is  drawn  from  the  chutes.  A  certain  amount  has  to  be  blasted 
again  as  it  packs.  The  ore  broken  in  the  stopes  before  the  recent  shut- 
down was  not  drawn  for  nearly  two  years  after  it  was  mined.  In  this 
case  a  considerable  amount  of  powder  had  to  be  used  to  loosen  the  packed 
ore,  on  which  account  only  a  few  of  the  square-set  timbers  could  be  saved. 
However,  if  the  ore  could  be  drawn  soon  after  being  broken,  the  amount 


JL 


Partly 
Waste 


fift 


Waste 


f---50- 


MM: 


zftft 


MM 


flft 


Waste 


MM 


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Plan  of  Sill  Floor 


TJ,e  Engineering  ^Mining  Journal 


The  Engineering  j  Mining  Journal 


FIG.   53. — Plan  Panel-core  stopes,  Veta  Grande        FIG.  54. — Vertical  section  Panel-core  stope,  Veta 
mine.  Grande  mine. 

of  the  powder  needed  would  be  less  and  a  large  percentage  of  the  timbers 
could  be  saved. 

After  all  the  ore  that  can  be  drawn  from  the  chutes  is  removed, 
there  will  still  remain  a  pyramid-like  mass  in  the  center  of  each  rectangle 
which  cannot  be  removed  in  this  way.  It  was  from  this  fact  that  the 
system  received  its  name.  The  pyramid  of  ore  is  later  drawn  by  driv- 
ing a  drift  into  the  center  of  the  block  and  with  a  raise  one  set  above 
the  sill  the  remaining  ore  is  drawn.  The  stoping  proper  does  not  com- 
mence until  16  ft.  above  the  level,  the  object  in  this  being  to  preserve 


OVERHAND    STOPING    WITH    SHRINKAGE  117 

the  level  drifts  with  16  ft.  of  solid  ore  above  the  rails  which  would  be 
mined  from  the  level  below. 

The  method  of  mining  the  block  of  ore  on  the  level  directly  below 
would  depend  entirely  upon  the  condition  of  the  waste  roof  to  which 
the  first  section  had  been  mined.  If  the  roof  were  treacherous  and 
unsafe,  it  would  be  caved  and  the  remaining  ore  could  be  mined  by  the 
slicing  system.  Fig.  53  shows  the  actual  method  of  blocking  out 
the  orebody.  As  shown,  chutes  are  put  in  every  other  set  with  no  two 
chutes  opposite  each  other,  as  this  would  obstruct  the  drift.  The  chutes 
are  merely  small  openings  cut  in  the  solid  ore  with  a  couple  of  chute 
jaws  and  a  door  attached  to  the  timbers.  Fig.  54  shows  a  section  across 
one  of  the  rectangles.  One  after  the  other  of  these  rectangular  blocks 
is  carried  all  the  way  up  to  the  waste  roof  and  the  drawing  of  the  ore 
does  not  commence  until  all  have  been  mined  out. 

This  system  takes  much  timber,  and  where  so  large  a  mass  of  broken 
ore  stands  before  being  drawn,  it  takes  considerable  labor  and  powder 
to  loosen  and  draw  it.  This  is  its  chief  disadvantage.  It  requires  solid 
ore  and  a  strong  roof  which  will  stand  over  the  core.  The  cost  for  the 
labor  and  timber  used  in  placing  ore  in  the  chutes  is  80  to  90  cents  a  ton. 


CHAPTER  X 
OVERHAND  STOPING  ON  WASTE  IN  THE  UNITED  STATES 

EXAMPLE    19. — SOUTH  RANGE  MINES,  HOUGHTON  COUNTY,  MICH 
(See  also  Example  14.) 

Dry-walled  Drifts.  Sub-vertical  Amygdaloid  Beds  with  Weak  Hanging 
Wall. — The  Quincy  mine  practice  will  first  be  reviewed,  as  the  prototype 
of  that  of  the  South  Range.  The  Quincy  amygdaloid,  proper,  is  overlaid 
by  a  shaly  seam  and,  to  avoid  this,  the  main  drifts  follow  the  footwall 
and  leave  the  lode  rock  overhead.  In  stoping  out  all  the  lode,  the  weak 
hanging  wall  must  be  supported,  and  this  has  been  done,  with  little  use 
of  timber,  by  a  system  of  dry-walling. 

The  present  South  Range  mining  system  was  -first  developed  at  the 
original  mine,  the  Baltic.  The  first  system  at  the  Baltic  was  that  of  its 


FIG.  55. — Dry-walled  tunnel,  Champion  mine. 

neighbor,  the  Atlantic  mine,  by  which  the  main  drifts  were  roofed  over 
with  long  stulls,  closely  lagged,  and  enough  of  the  ore,  broken  in  the 
overhand  stope  above,  was  left  on  these  stulls  to  hold  the  drillmen  close 
to  the  stoping  face.  Enough  room  was  left  between  the  hanging  wall 
and  the  stulls  (which  were  inclined  at  about  70  deg.,  with  one  end  on  the 
sill  floor,  and  the  other  in  a  hitch  on  the  hanging  wall),  for  the  track, 
where  cars  were  filled  from  chute-gates  above.  When  the  stope  was 
completed  to  the  15-ft.  longitudinal  rib,  to  the  left  under  the  next  higher 

118 


OVERHAND    STOPING    ON    WASTE   IN   THE   UNITED    STATES 


119 


level,  the  withdrawal  of  its  content  of  broken  ore  was  begun.  During 
this  last  process,  stulls  were  placed  between  the  walls  by  the  timbermen 
at  dangerous  places;  but,  nevertheless,  considerable  hanging  wall  would 
peel  off  and  contaminate  the  ore.  The  net  result  was  that  20  per  cent, 
of  the  ore  was  waste  and,  as  all  sorting  had  to  be  done  at  the  surface,  this 
caused  decreased  hoisting  capacity  for  mill-ore.  Also,  the  stull  system 
required  a  regular  width  of  stope  and,  in  the  irregular  Baltic  lode,  this 
meant  either  an  unnecessary  breaking  down  of  waste  or  the  missing  of 
bulges  of  ore. 

The  Champion. — This  mine  was  first  opened  by  "  arching/'  by  which 
a  stone  arch  8  ft.  thick  was  left  above  each  main  drift,  with  chutes  cut 
through  it  at  25-ft.  intervals.  This  arch  corresponded  to  the  lagged- 
stull  roof  of  the  Altantic  system;  and  enough  of  the  broken  ore  was  left 


FIG.  56. — Stoping  at  Champion  mine. 


in  it  to  hold  the  drillers  up  to  the  face  of  the  overhand  stope  above. 
" Arching'  had  the  same  disadvantages  at  the  Champion  as  "long- 
stulling"  had  at  the  Baltic,  and  has  now  been  superseded  by  "  dry- 
walling"  as  adapted  by  the  Baltic  in  1900.  The  Champion  mine's 
"  dry- walling "  will  be  described  as  typical  of  the  present  South  Range 
system. 

The  main  levels  are  driven  100  ft.  apart  on  the  70  deg.  footwall,  and 
are  cut  out  the  whole  width  of  the  ore,  whether  10  or  60  ft.  The  rock  is 
sorted  where  broken,  the  ore  being  hoisted,  and  the  waste  used  for  walling, 
as  shown  in  Figs.  57  and  58.'  The  side  walls  of  drift  D.  (Fig.  56)  are  laid 
dry  and  4  ft.  thick  at  the  base.  The  walls  are  topped  with  2-in.  plank, 
on  which  are  laid  (at  5-ft.  centers)  the  unbarked,  round  caps  C,  of  8  to 
10-in.  diameter,  which  support  the  lagging  L  of  3-in.  poles.  The  drift  is 
located  near  the  footwall,  but  alloys  space  for  the  ore  chute  N,  set  every 
25  to  60  ft.  along  the  drift,  with  hinged,  steel  troughs  for  gates. 

As  soon  as  the  drift-lining  is  well  advanced  and  well  backed  by  waste 
G,  the  overhand  benches  are  begun  behind  it.  In  Fig.  56  two  machines 
are  stoping,  while  the  sorters  are  handling  the  freshly  broken  lode, 


120 


MINING    WITHOUT    TIMBER 


throwing  the  ore  into  chute  N,  and  piling  the  waste  at  G  and  F.  The 
chute  N  is  kept  just  above  the  stowing,  and  is  built  (4  ft.  inside  diameter) 
of  waste  rubble,  which  has  proved  superior  to  the  poles  and  old  railroad 
ties  formerly  used.  Should  the  waste  prove  insufficient  for  filling  the 
stope,  some  can  be  blasted  from  the  walls,  or  a  raise  can  be  put  through 
to  the  next  level,  and  waste  run  down  from  the  old  filling  above.  When 
the  exhausted  upper  level  is  reached,  the  floor-pillar  can  be  extracted  by 
caving,  if  the  level  need  no  longer  be  kept  open  for  tramming.  A  small 
self -dumping  car  is  often  used  in  a  stope  to  facilitate  the  stowage  of  waste. 

By  this  system,  a  total  of  1100  to  1200  men  (above  and  below  ground) 
produce  2500  to  2800  tons  in  two  shifts.  Owing  to  the  irregular  stope- 
width,  miners  work  by  day's  pay,  earning  about  12.50,  while  muckers 
get  $52  to  $54  a  month.  Only  development  is  let  on  contract. 

The  advantages  of  the  South  Range  system  are,  the  complete  exhaus- 
tion of  the  ore,  the  saving  in  timber,  and  the  decreased  use  of  shafts  for 
lowering  timber  and  hoisting  waste,  all  accompanied  by  safety  and  good 
ventilation.  A  little  copper  is  lost  in  the  stored  waste  and  many  men 
are  needed  for  dry-walling;  though  with  the  large  output,  this  only 
means  an  expense  of  about  8  cents  per  ton  of  ore  hoisted. 

TKAMMING 

It  is  only  in  the  South  Range  mines,  with  their  steep  footwalls,  that 
the  ore  broken  in  the  stopes  can  all  be  drawn  direct  from  chutes;  in  the 
rest  of  the  district,  it  is  either  shoveled  off  a  sollar,  placed  on  the  drift- 
floor,  or  on  a  platform  at  one  side,  so  that  the  car  can  pass  to  stopes 


12  Loose  Wheel- 
Fio.  57. — Car  at  Champion  mine. 


-  — 3-Gauge  - 
Fixed  Wheel' 


beyond.  In  capacity,  the  cars  vary  from  2  to  3  tons;  the  latter  size  is 
unusually  large  for  man-power,  but  there 'seems  to  be  no  difficulty  in 
handling  it  with  two  men  on  a  track  of  3-ft.  gauge  at  the  Champion  mine; 
but  for  longer  hauls  electric  trains  are  cheaper,  and  are  used  at  the 
Quincy  with  3-ton  locomotives. 

In  such  cars  as  the  Wolverine  and  the  Champion  (Fig.  57),  the  ideal 
design  for  capacity  and  ease  of  loading  seems  to  have  been  attained. 
The  car  body  is  placed  low  (just  above  the  rim  of  the  12-in.  wheels),  and 
is  only  2  ft.  high,  for  easy  loading  from  the  track  level,  while  capacity 


OVERHAND    STOPING    ON    WASTE    IN    THE    UNITED    STATES  121 

is  had  by  the  extreme  length  of  7  ft.  for  the  first  and  9  ft.  for  the  second. 
In  order  to  get  a  low  car-body,  a  truck  with  a  turn-plate  is  barred;  so 
dumping  is  achieved  for  the  Wolverine  car  by  rotating  its  body  around 
the  front  axle,  while  the  larger  Champion  car  must  be  run  on  a  tipple  to 
be  emptied.  The  Wolverine  car  has  loose  wheels,  but  the  Champion  has 
one  loose  and  one  tight  wheel  on  each  axle,  which  gives  better  lubrication 
and  less  wear. 

The  copper  region  has  been  fortunate  in  having  a  bedded  formation 
of  great  continuity,  uniformity  and  strength;  if  it  had  been  much  faulted 
or  generally  brittle,  the  system  of  great  open  stopes  could  not  have  been 
pursued,  and  consequently  the  poorer  lodes  could  not  have  been  profit- 
ably worked.  Nevertheless,  there  is  a  limit  to  even  the  stability  of  these 
formations,  and  this  fact  has  been  forcibly  emphasized  by  recent  events. 

For  many  years  some  of  the  older  mines,  notably  the  Quincy,  had 
been  bothered  with  subterranean  disturbances  due  to  settling  of  the 
hanging  wall  in  the  old  stopes;  but  it  was  only  recently  that  a  calamity 
resulted.  In  May,  1906,  the  main  workings  of  the  Atlantic  Co. 
on  the  Ash-bed  lode  collapsed,  and  subsequent  movements  soon  rendered 
the  5000-ft.  shafts  useless  for  further  hoisting.  The  company  has  not 
tried  to  reopen  the  shafts,  but  has  been  fortunate  in  finding  a  new  mine 
on  its  holdings  along  the  Baltic  lode. 

One  explanation  of  such  general  caves  is  that  the  adherence  of  the 
hanging  wall  to  the  stope-pillars  is  so  lessened  by  the  great  area  of  nearly 
continuous  excavation  that  it  begins  to  slip  along  its  sustaining  pillars, 
and  thereby  so  crushes  and  distorts  them  that  they  no  longer  offer 
sufficient  support  to  the  superincumbent  weight.  The  hanging  wall  of 
an  isolated  stope  can  be  considered  as  a  beam  fixed  around  its  circum- 
ferential supports;  but  when  many  stopes  are  connected,  the  hanging 
wall  is  then  only  like  a  beam  resting  on  its  supports,  and  has  consequently 
diminished  stability.  Also,  a  pillar,  that  is  strong  enough  to  sustain  the 
roof  of  the  single  original  stope,  is  not  necessarily  able  to  sustain  the 
increased  strain  of  an  added  line  of  contiguous  stopes. 

EXAMPLE  20. — MINNESOTA  MINE,   SOUDAN,   VERMILION  IRON  RANGE, 

MINNESOTA 

(See  also  Example  36.) 

Overhand  Sloping  on  Waste  with  Filling  from  a  Descending  Hangivall. 
Sub-vertical  Wide  Vein  with  Weak  Hangwall. — These  iron  ore  deposits 
occur  in  lenses  200  ft.  to  1000  ft.  long  and  5  to  80  ft.  wide,  and  stand  at 
an  angle  of  65  to  75  deg.,  with  a  vertical  height  of  250  ft.  to  500  ft.,  other 
lenses  occurring  below.  A  number  of  the  deposits  were  first  worked  as  open 
pits,  which  in  some  cases  were  carried  to  depths  of  150  ft.,  when,  owing 
to  the  weakness  of  the  walls,  underground  mining  was  adopted.  While 


122  MINING    WITHOUT   TIMBER 

the  ore  was  being  removed  from  the'open  pit,  shafts  were  in  several  ins- 
tances sunk  into  the  foot-wall,  the  intention  being  to  mine  the  ore  with 
breast-stopes  of  an  approximate  height  of  20  ft.,  followed  by  underhand 
stopes  of  the  same  hight,  leaving  pillars  between  of  the  necessary  thick- 
ness to  support  the  walls.  As  work  progressed,  however,  it  was  found 
that  the  chlorite  walls  were  too  weak  to  permit  the  working  of  breast- 
stopes  20  ft.  high,  there  being  frequent  heavy  falls  of  ground  from  the 
hanging  wall,  and  sometimes  from  the  foot.  The  plan  of  following 
breast-stopes  with  underhand  stopes  was  therefore  abandoned,  for  by 
working  breast-stopes  only,  but  little  more  than  one-half  of  the  ore  could 
be  removed,  and  that  only  at  an  excessive  cost,  the  ore  being  one  of  the 
hardest  known,  to  drill. 

The  Rand  3  1/8-in.  piston  type  is  used,  with  60  Ib.  of  air,  which  drills 
6  ft.  per  shift  as  a  yearly  average,  but  in  certain  places  will  only  make  1  ft.  in 
10  hours.  During  a  shift's  work  each  drill  dulls  45  to  50  bits,  and  even 
then  powder  must  be  used  as  an  aid  by  exploding  a  half  stick  of  50  per  cent, 
in  the  hole  for  every  few  inches  of  advance,  to  enlarge  the  bottom  and 
prevent  the  bit  sticking.  Often  10  sticks  of  powder  are  used  in  boring 
a  hole  and,  in  addition,  there  are  5  to  10  sticks  more  necessary  to 
chamber  the  bottom  for  the  breaking  charge  of  20  to  50  sticks. 
Fuse  and  cap  firing  is  in  vogue.  The  holes  are  6  ft.  to  10  ft.  deep. 
For  a  time  diamond  drilling  was  employed  to  bore  20-ft.  to  40-ft.  holes 
for  breaking  ground,  but  the  present  high  price  of  diamonds  has  made 
this  method  unprofitable. 

The  greenstone  walls  are  easy  to  drill,  and  as  much  as  130  ft.  per 
month  has  been  drifted  in  them  with  two  drill  shifts  daily.  But  in  the 
jaspilite  " horses"  encountered  the  same  men  could  only  make  one-third 
this  distance.  In  the  sinking  of  the  main  shaft  below  the  1,150-ft. 
level,  with  three  machines,  the  monthly  advance  in  jaspilite  with  three 
eight-hour  shifts  was  only  12  ft.  to  18  ft. 

Mining. — The  present  system  here  may  be  called  hangwall  filling. 
To  develop  it  an  incline  was  sunk  in  the  foot  wall,  and  from  it,  at  about 
100-ft.  vertical  intervals,  were  driven  crosscuts  in  the  vein.  The  ore 
is  then  attacked  in  all  directions  by  overhand  stoping  until  the  excava- 
tion is  16  ft.  by  20  ft.  high  the  whole  length  and  width  of  the  ore  body 
(see  Fig.  58). 

Next  drift  sets  d  are  set  up  and  spiked  together  the  whole  length  of 
the  stope  with  9-ft.  caps  and  posts,  but  no  sills.  At  25-ft.  intervals  on 
the  footwall  side  chutes  c,  5  ft.  square  inside,  are  built  along  the  drift 
sets,  and  opposite  every  third  chute  is  placed  a  similarly  constructed 
manway.  The  chutes  rest  on  the  floor,  so  they  must  be  filled  with 
waste  up  to  their  false  bottom  of  rails,  which  is  high  enough  to  deliver 
through  a  hinged,  steel  spout  into  the  ore  car.  Above  the  false  bottoms 
the  chutes  are  lined  with  2-in.  planks  placed  vertically. 


OVEKHAND    STOPINQ    ON    WASTE    IN   THE    UNITED    STATES 


123 


Meanwhile,  waste  raises  w  have  been  driven  at  75-ft.  intervals  in  the 
foot  wall,  just  under  the  ore,  from  the  stope  to  the  level  above.  These 
waste  raises  extend  from  the  open  pit  to  all  the  opened  levels  and  are 
cribbed  up  in  front  so  that  waste  can  be  thrown  at  any  point  by  removing 
some  cribbing  and  leaving  the  part  below  full,  or  making  a  temporary 
false  bottom  of  wood.  Raises  /  are  also  put  through  between  levels 
for  use  as  extra  manways  in  case  of  fire. 

When  the  timbering  and  waste  raises  are  completed  to  the  lowest 
floor  the  filling  of  excavation  E  is  begun.  The  waste  for  the  first 
level  below  the  open  cut  is  obtained  from  the  hangwall  either  by  its 


Cross  Sec.^~ 
FIG.  58. — iStoping  at  Minnesota  mine. 

natural  caving  or  by  blasting.  If  waste  for  a  lower  level  is  not  needed 
till  the  first  level  is  all  stoped,  much  of  it  can  be  drawn  from  the  filling 
of  the  latter,  as  in  winter  the  overlying  waste  on  pillar  P  is  frozen  hard 
enough  to  prevent  its  breaking  through. 

Wherever  obtained,  the  waste  is  drawn  down  from  raises  w  and 
handled  in  wheelbarrows  to  fill  the  whole  stope  to  a  depth  of  14  ft., 
the  drift  sets  being  lagged  and  the  chutes  and  manways  being  extended 
to  the  same  height.  The  drills  are  then  set  up  on  the  filling  and  a  14-ft. 
to  16-ft.  cut  made  to  the  back  to  start  a  new  breast  stope,  for  which 
the  drill  is  to  be  worked  from  a  scaffold.  Filling  can  begin  behind  the 
breaking,  the  ore  being  thrown  into  the  chutes  ahead  and  the  waste 
brought  from  the  raise  ID  behind  and  piled  to  within  6  ft.  of  the  new  back, 
while  the  chutes  and  manways  are  correspondingly  heightened. 


124  MINING    WITHOUT   TIMBER 

This  overhead  slicing  is  continued  till  the  next  level  is  nearly  reached, 
under  which  a  pillar  P  is  left  from  6  ft.  to  10  ft.  high,  according  to  ground. 
During  the  ascent  any  weak  parts  of  the  back  are  supported  by  cribs 
on  top  the  filling,  which  can  be  removed  on  stoping  the  next  slice.  When 
all  ore  above  is  removed  and  the  filling  in  the  open  cut  has  settled  on 
pillar  P,  it  can  be  removed  by  beginning  at  each  end  of  the  ore  body  and 
holing  through  by  blasting.  The  balance  of  the  pillar  is  then  cut  off 
while  retreating,  and  though  the  filling  follows  the  broken  ore  down, 
the  latter  comes  first,  and  can  thus  be  recovered  and  thrown  into  the 
chutes.  During  the  recovery  of  P  it  is  liable  to  cave  from  the  pressure 
of  the  waste  above,  but  as  P  is  always  supported  by  a  number  of  cribs 
on  the  filling,  and  gives  plenty  of  advance  warning  of  independing  disaster, 
the  men  are  in  little  danger. 

In  case  pillar  P  does  cave  it  is  allowed  to  settle  and  its  ores  recovered 
by  drifting.  For  this  it  is  necessary  to  drive  spiling  ahead,  which  are 
of  sharpened  poles,  4  in.  by  16  ft.,  driven  over  three-quarter  drift  sets 
placed  3  ft.  apart.  To  advance  the  spiling  it  often  requires  considerable 
blasting  of  the  many  boulders  encountered. 

Application  of  Hangwall  Filling  System. — The  following  are  favor- 
able conditions:  Highly  inclined,  hard,  wide  veins,  which  will  stand 
without  timbers  when  excavated  across  their  whole  width,  but  whose 
hangwalls  are  weak  and  friable. 

It  allows  good  ventilation  and  requires  little  timber  and  no  hoisting 
of  waste.  Several  levels  can  be  worked  simultaneously  and,  develop- 
ment being  confined  to  the  soft  walls,  the  hard  ore  can  be  broken  in  wide 
stopes  with  minimum  expense  for  drilling  and  explosives,  as  most  of  the 
filling  comes  from  the  natural  shelling  of  the  hangwall,  much  of  which 
can  be  reused,  there  is  no  expense  for  mining  or  freight,  the  only  cost 
being  in  keeping  the  waste  raises  clear  and  in  wheeling  and  stowing  the 
filling  in  place.  For  narrower  veins  it  is  sometimes  cheaper  to  break 
part  of  the  filling  from  the  hangwall  of  the  stope. 

EXAMPLE  21. — SUPERIOR  AND  BOSTON  COPPER  MINE,  GLOBE  DISTRICT, 

ARIZONA 

(See  also  Examples  32  and  42.) 

Sub-vertical  Wide  Vein  with  Weak  Hanging-wall;  Rill  Chutes. — The 
present  mine  production  is  made  mainly  from  the  stopes  of  the  550-foot 
level.  Here  the  vein  dips  about  58  deg.  and  varies  in  thickness  from 
7  to  15  ft.,  but  averages  from  9  to  10  ft.  The  foot-wall  is  hard  and 
smooth,  but  on  the  hanging  wall  the  ore  is  " frozen"  and  there  is  no  de- 
fined wall.  This  method  of  mining  was  devised  by  Supt.  John  D. 
Wanviz,  and  is  similar  to  one  in  use  at  Zaruma,  Ecuador.* 

*Trans.  A.  I.  M.  E.,  Vol.  XXX,  p.  248. 


OVERHAND    STOPING    ON    WASTE    IN   THE    UNITED    STATES 


125 


Until  recently  the  ore  has  been  mined  by  the  common  square-set 
system,  but  in  the  few  months  since  the  first  trial  of  the  new  system, 
both  the  timber  and  labor,  not  to  mention  loss  of  ore  fines,  have  been 
materially  reduced,  and  the  reduction  in  costs  has  been  60  per  cent. 

Reference  to  Fig.  59  will  make  clear  the  following  data: 

Drifts  5x7  ft.  are  carried  100  ft.  apart  vertically  or  about  117  ft. 
along  the  58-deg.  dip  of  the  vein. 

Two-compartment  raises  are  put  up  at  100-ft.  intervals  to  connect  the 
drifts;  the  chute  compartment  is  4x4  ft.,  and  the  ladderway  3x4  ft. 
These  raises  are  on  the  foot-wall  and  are  timbered  only  by  two  lines  of 
stulls  with  head-boards  and  one  set  of  lagging. 

The  division  between  the  two  compartments  is  formed  by  the  first 
line  of  stulls  which  is  plank  lagged.  The  second  line  of  stulls  is  carried 


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FIG.  59. — Sloping  at  Superior  and  Boston  mine. 

on  the  other  side  of  the  ladderway  compartment,  but  this  is  unlagged. 
The  outer  wall  of  the  chute  is  thus  constituted  by  a  rock  wall  unstulled 
and  unlagged  and  the  outer  wall  of  the  ladderway  is  likewise  constituted 
by  a  rock  wall,  but  this  is  stulled.  The  main  drifts  are  all  timbered  with 
drift  sets  spaced  on  5-foot  centers  and  heavily  lagged.  Chutes 
4  ft.  wide  are  spaced  on  15-ft.  centers,  the  chute  gates  all  being  placed 
at  once,  but  the  chutes  themselves  are  put  up  one  by  one  as  stoping 
proceeds  above.  The  chute  mouths  afford  one  of  the  ways  of  access  to 
a  stope  as  it  progresses,  since  there  is  always  an  irregular  space  1  to  3  ft. 
high  between  the  ore  back  and  the  top  of  the  drift  lagging. 

Stoping  is  started  at  the  lower  corners  of  two  adjacent  blocks  formed 
by  the  intersection  of  a  raise  with  a  drift.  These  two  blocks  are  simul- 
taneously stoped  up  along  the  raise  and  retreating  from  it  horizontally 


126  MINING    WITHOUT   TIMBER 

as  shown.  At  such  a  corner  the  back  is  drilled  with  stoping  drills  and 
broken  down  on  to  the  drift  lagging  for  a  distance  of  8  to  10  ft.  The 
broken  ore  is  at  first  discharged  into  mine  cars  standing  in  the  drift 
below  by  pulling  out  the  lagging  and  letting  it  run  down. 

The  first  chute,  1,  is  then  placed.  This  chute  is  formed  by  carrying 
up  two  lines  of  stulls  4  ft.  apart  along  the  strike  of  the  vein  and  lagging 
them  on  the  outside  with  plank. 

The  stulls  are  placed  inside  the  plank  lagging  instead  of  outside, 
as  in  ordinary  practice.  The  reason  for  this  is  found  in  the  fact  that 
when  the  chutes  are  abandoned  they  are  not  filled  with  waste,  so  that 
empty  they  have  to  stand  the  pressure  of  the  waste  filling  on  all  sides. 

The  entire  width  of  vein  from  foot  to  hanging  wall  is  broken  out 
between  the  rows  of  stulls.  With  the  first  chutes  in  place,  the  timber  in 
the  lower  part  of  the  raise  b  is  taken  out  and  waste  c  for  filling  dumped 
down  the  chute  from  the  drift  above.  This  waste  assumes  a  natural 
slope  of  about  37  deg.  from  the  horizontal,  and  it  accordingly  spreads 
out  from  the  foot  of  the  raise  into  the  stopes  on  each  side  of  it  at  that 
angle,  running  farther  and  farther  horizontally  into  the  stopes  as  they  are 
advanced  and  the  pile  of  waste  increases  in  height.  In  the  first  stage  of 
the  work,  however,  advancing  toward  chute  No.  1,  the  waste  is  dumped 
down  till  the  foot  of  its  slope  almost  reaches  the  chute.  Dumping  waste 
then  ceases  for  the  time  being.  With  sloping  floor  of  the  waste  now 
brought  within  5  to  7  ft.  of  the  ore  back,  a  sloping  wooden  floor  d  is  laid 
down  on  top  of  the  waste  extending  from  the  top  of  the  chute  to  the 
manway  of  the  raise.  Its  purpose  is  to  receive  the  ore  broken  from  the 
back  and  discharge  it  into  the  chute,  thus  serving  the  double  purpose  of 
separating  the  ore  from  the  waste  and  eliminating  all  shoveling  and 
tramming  in  the  slope.  It  will  be  evident  that  since  the  waste  will  run 
at  an  angle  of  37  deg.,  the  ore  will  certainly  run  on  the  wooden  planks 
which  are  laid  over  the  waste  on  the  same  angle. 

The  width  of  the  plank  floor  d  is  everywhere  made  that  of  the  vein 
from  wall  to  wall,  and  its  length,  of  course,  increases  up  to  a  maximum 
of  about  60  ft.  in  extending  from  the  center  chute  of  the  block  up  to 
the  raises  on  either  side  of  it.  The  floor  consists  of  2-in.  plank  and  15-ft. 
lengths  nailed  together  by  cross-pieces  to  form  sections  about  2  ft. 
wide.  In  laying  down  these  sections  over  the  waste,  the  ends  are  made 
to  form  a  butt  joint  in  order  to  insure  a  clean  run-off  of  the  ore.  With 
the  floor  laid  down,  the  stope  drills  are  put  at  work  to  break  down 
the  new  back,  stulls  and  planks  being  placed  where  necessary  to  afford 
a  good  secure  footing  next  to  the  back.  When  drilling  operations  are 
completed,  the  stulls  are  pulled,  thus  recovering  them  for  further  use. 
The  chute  at  the  foot  of  the  sloping  floor  is  always  kept  nearly  full  of 
ore  so  that  wear  upon  it  may  be  kept  at  a  minimum.  After  drilling  and 
shooting  the  back,  the  ore  is  at  once  drawn  down  to  the  top  of  the  chute, 


OVERHAND    STOPING    ON    WASTE   IN   THE    UNITED    STATES  127 

thus  clearing  the  floor.  A  number  of  stulls  are  next  put  in  as  close  to  the 
back  as  possible  and  the  sections  of  the  platform  are  then  raised  and  one 
of  each  placed  see-saw  fashion  over  the  stulls  so  that  the  weight  of  a 
larger  end  of  a  given  section  causes  the  smaller  end  to  press  up  against 
the  ore  back  and  thus  be  firmly  held.  With  the  platform  sections  thus 
disposed  conveniently  at  hand  for  their  next  period  of  use,  the  top  of  the 
chute  is  now  built  up  and  more  waste  dumped  down  the  raise  till  the  foot 
of  its  slope  has  nearly  reached  the  level  of  the  chute  top.  The  laying  of 
the  floor  and  the  mining  of  another  diagonal  slice  now  proceeds  as  before. 

A  given  chute  can  be  used  at  the  foot  of  the  sloping  platform  only 
till  the  angle  of  slope  carries  down  the  ore  in  line  with  the  center  of  its 
top.     After  that  the  top  is  lagged  over,  the  chute  emptied  of  its  ore 
but    not    filled    with    waste,    and    it    is   abandoned.     Fig.   59  shows 
chute   No.   1  just   abandoned   and    buried  in  waste  with  the  use  of 
chute  No.  2  just  begun.     The  illustration  also  makes  is  evident  that  of 
the  seven  chutes  raised  in  a  100-ft.  block,  all  but  the  one  in  the  center 
are  limited  to  a  height  of  about  10  ft.,  as  at  that  height  the  angle  of  waste 
slope  carries  the  flow  past  to  the  bottom  of  the  next  adjacent  chute, 
which  is  likewise  built  up.     By  working  from  the  two  opposite  lower 
corners  of  the  block  simultaneously,  and  thus  advancing  the  diagonal 
slices  toward  the  center  line  of  the  block,  they  finally  intersect  at  the 
bottom  of  the  centrally  placed  chute  No.  4.     From  this  point  on,  chute 
No.  4  receives  all  the  ore  of  the  block.     It  is  built  up  from  time  to  time 
just  as  the  other  chutes  were,  with  the  exception  that  it  is  more  heavily 
and    completely   lagged   by   placing   2-in.    lagging    outside   the   stulls 
and  4-in.  plank  lagging  inside  the  stulls  which  bear  the  wear  of  the 
chutes.     Its  height,  as  shown  in  dotted  lines,   is  limited  only  by  the 
floor  of    the  drift  above.     It  will  likewise  be  evident  from  the  illus- 
tration that  when  the  top  of  the  waste  filling  has  reached  the  approxi- 
mate position  of  the  dotted  line  /,  g,  h,  the  drift  e  must  be  abandoned 
for  through  traffic,  or  its  floor  supported  on  stulls.     From  this  point  on, 
the  raises  cannot  be  utilized  for  throwing  down  filling,  so  that  waste  is 
dumped  from  points  along  the  drift  itself  advancing  toward  the  center 
line  of  the  block  till  the  central  chute  reaches  the  drift  and  the  block 
is  entirely  mined  and  filled. 

During  the  mining  of  the  lower  portion  of  the  block,  access  is  had  to 
the  stope  from  the  manways  of  the  raises  or  from  the  chute  mouths  in 
the  lower  drift  roof.  After  the  diagonal  slices  have  connected  at  the 
central  chute,  No.  4,  however,  access  is  had  only  from  the  drift  e  above 
by  descending  through  the  raise  manways  into  the  stope.  No  objection 
to  this  limitation  of  accessibility  has  yet  been  found  in  several  months  of 
work.  If  found  desirable,  however,  the  floor  on  one  side  or  the  other 
could  always  be  kept  down  and  over  it  the  men  could  pass  in  and  out  to 
avoid  danger  of  starting  the  loose  waste. 


128 


MINING    WITHOUT    TIMBER 


From  the  loose  nature  of  the  hanging  wall,  it  will  be  evident  that 
working  the  mine  by  "shrinkage  stoping"  would  be  inadvisable,  because 
caves  from  the  hanging  wall  would  seriously  dilute  the  ore  with  waste. 
The  system  adopted,  as  described  above,  therefore,  seems  the  next 
best  method  of  reducing  the  use  of  timber  to  a  minimum,  eliminating 
tramming  and  shoveling  in  stopes,  and  recovering  all  the  fine  ore.  The 
system  has  the  advantage  over  "  shrinkage  stoping"  of  making  all  the 
ore  broken  available  at  once  instead  of  only  about  25  to  40  per  cent, 
until  such  time  as  the  stope  is  finished.  Another  advantage  over 

"shrinkage  st oping"  is  that  the 
drift  timber  sets  and  chutes  once 
placed  require  no  reinforcement 
during  the  drawing  of  ore  as  is 
frequently  the  case  in  drawing  a 
stope  completed  by  the  shrinkage 
method. 

EXAMPLE  22. — METCALF  MINE, 
GRAHAM  COUNTY,  ARIZ. 

(See  also  Examples  29,  30  and  40.) 

Irregular  Lenses  in  Porphyry; 
Auxiliary  Milling  and  Square  Set- 
ting.— At  this  mine,  bodies  of  oxi- 
dized ore  form  a  conspicuous  out- 
crop on  the  hilltop,  (Fig.  60)  and 
were  the  source  of  the  high-grade 
ore  first  mined.  The  hill  is  a  mass 
of  granite  porphyry,  capped  by 
the  lower  members  of  the  sedi- 
mentary rock  series.  Development 
has  demonstrated  the  existence  of 
four  parallel  vein  systems  or  stock- 
works  in  the  granite  porphyry, 
along  which  oreshoots  of  varied 
magnitude  are  found.  The  vein 
systems  have  suffered  severe  cross  faulting,  subsequent  to  the  forma- 
tion of  the  primary  ore,  but  prior  to  the  surface  enrichment  that  has 
occurred.  The  surface  presents  a  chaotic  mass  of  blocks  of  quartzite, 
shale  and  granitized  limestone  lying  on  and  in  other  places  completely 
imbedded  in  the  intruded  porphyry. 

The  oreshoots  are  generally  found  at  the  junctions  of  the  cross  faulting 
with  the  vein  systems.  *  Although,  as  before  mentioned,  some  of  the 
oreshoots  outcropped,  the  majority  of  them  are  found  beneath  an  over- 


FIG.  60. — Metcalf  mine  and  ore  bins. 


OVEEHAND    STOPING    ON    WASTE    IN   THE   UNITED   STATES 


129 


burden  of  barren  rock  of  varying  thickness.  The  horizon  on  which  these 
shoots  occur  is  variable  and  their  discovery  necessitates  extensive  pros- 
pecting from  levels  of  not  more  than  40  to  50  ft.  apart. 

Where  an  oreshoot  has  been  proved  in  depth  and  is  covered  by  a 
heavy  overburden  of  waste,  underground  mining  is  employed.  The 
ground  in  most  cases  stands  well  without  timbering;  stopes  up  to  75  ft. 
in  width  having  been  worked  without  difficulty.  In  ore  of  moderate 
hardness,  the  mode  of  working  is  as  follows: 

On  the  lowest  level  on  which  the  ore  is  exposed,  the  oreshoot  is  opened 
to  its  full  width  and  length.  When  the  shape  of  the  orebody  has  been 
determined,  raises  are  made  from  the  roof  of  the  stope  to  the  surface, 
the  number  depending  on  the  dimensions  of  the  stope.  The  ore  is  then 
mined  to  a  height  of  20  ft.  above -the  level  and  a  timbered  roadway  with 
the  necessary  chutes  and  ladderways  erected.  The  overburden  of  waste 
is  now  milled  down  the  raises  and  leveled  off,  filling  the  stope  and 
forming  a  compact  working  floor,  15.  ft.  above  the  level. 


FIG.  61. —  Underground  milling  at  Metcalf  mine. 

The  ore  is  afterward  broken  by  overhand  stoping  in  ascending  slices 
15  to  25  ft.  high;  depending  on  the  condition  of  the  roof.  As  each  slice 
is  removed,  the  overburden  is  milled  down  for  filling  material  and  to 
provide  the  next  working  floor.  When  all  the  overburden  has  been 
utilized  in  this  manner,  the  back  of  ore  remaining  is  gained  by  open-cut 
or  milling  method.  The  cost  of  breaking  the  ore  in  stoping  is  necessarily 
higher  than  in  open-cut  work,  but  the  more  economical  removal  of  over- 
burden compensates  for  this  increase. 

When  the  ore  is  hard  and  stands  exceptionally  well  the  method 
of  extraction  is  shown  in  Fig.  61.  A  raise  is  made  to  the  top  of  the  ore 


130  MINING    WITHOUT    TIMBER 

and  extended  to  the  surface  or  to  an  upper  working  for  air.  At  a  height 
of  50  ft.  above  the  level,  mining  is  started  from  the  raise  outward,  the 
floor  being  always  left  sloping  so  that  the  ore  will  run  directly  into  the 
chute.  When  the  extremities  of  the  ore  have  been  reached  or  the 
roof  is  as  wide  as  will  stand  with  safety,  the  bench  forming  the  floor  of 
the  stope  is  now  mined.  Deep  holes,  charged  with  black  powder,  are 
used  and  the  ore  is  broken  as  freely  as  in  open-cut  work.  The  slower 
action  of  the  black  powder  does  not  jar  and  weaken  the  roof  of  the 
stope  to  the  same  degree  as  the  rapid  action  of  dynamite.. 

When  the  first  bench,  50  ft.  in  height,  has  been  worked  out,  the  chute 
and  ladderway  are  timbered  to  within  5  ft.  of  the  roof  and  the  stope  is 
filled  with  waste  from  the  surface;  the  filling  material  is  leveled  and  the 
next  block  of  ore  above  is  attacked  in  a  similar  manner. 

STOPING  WITH  SQUARE  SETS 

In  some  parts  of  the  mine,  the  ore  is  too  soft  and  friable  to  permit 
of  any  system  of  stoping  without  the  use  of  timbers.  Ordinary  square- 
set  timbering  is  employed  in  such  cases  to  support  the  roof  and  walls. 
The  stopes  are  kept  full  of  waste  to  within  one  set  of  the  back  of  ore, 
timbered  chutes  and  ladderways  alone  being  left  open.  The  waste  is 
obtained  from  the  surface  workings  and  is  distributed  from  a  small 
chute  placed  in  one  corner  of  the  stope.  This  position  of  the  chute 
is  advisable  as  by  commencing  the  mining  of  each  floor  from  this 
point,  the  filling  can  be  kept  close  to  the  working  breast  without  inter- 
fering with  the  shovelers. 

Every  floor  is  worked  as  rapidly  as  the  working  faces  allow,  to  avoid 
excessive  weight  settling  on  the  timbers.  The  timbering  is  arched  to 
compensate  for  the  sinking  of  the  sets  in  the  center  of  the  stope  due  to 
the  greater  weight  of  roof.  This  is  accomplished  by  introducing  one 
floor  of  sets,  the  posts  of  which  step  upward  2  in.  to  the  center  of  the 
stope,  and  descending  in  like  manner  to.  the  opposite  wall.  Should  this 
arch  effect  at  any  time  be  lost  by  excessive  weight,  causing  the  timbers 
to  settle,  or  by  the  loosening  of  the  side  blocking,  another  floor  of  special 
posts  is  put  in  to  restore  it.  Further  details  are  given  in  example  23. 

EXAMPLE   23. — COPPER  QUEEN  MINES,  BISBEE,  ARIZ. 
(See  also  Example  12.) 

Irregular  Clayey  Lenses  in  Limestone;  Panel  System  and  Square 
Sets. — Of  the  gangue  minerals,  the  silica  is  not  vein  quartz,  but  fine 
grained  aggregates,  and  in  some  stopes  it  runs  like  pulverulent  sand; 
while  the  silicates  occur  interspersed  throughout  the  altered  limestones 
around  the  ore.  Clay  is  prevalent,  it  may  be  white,  gray,  or  yellow 


OVERHAND    STOPING    ON    WASTE    IN    THE    UNITED    STATES  131 

and  red  from  limonite  admixture,  and  it  forms  the  chief  gangue  of  the 
oxidized  stopes  and  makes  their  support  difficult. 

The  ore  horizon  is  the  Escabrosa  limestone  with  occasional  offshoots 
into  the  Naco  above.  The  ore  occurs  in  great  tables  and  lenses  that 
follow  the  bedding  planes  of  the  beds.  Within  the  limestone  the  bodies 
pinch  and  swell  and  many  of  them,  particularly  in  the  central  belt,  are 
connected  by  seams  and  pipes. 

Definite  walls  are  exceptional;  the  oxidized  ores  grade  into  clay  and 
the  sulphides  often  into  oxides  within  a  hard  limestone  casing.  A  stope 
outline  depends  upon  the  price  of  copper  as  compared  with  the  cost  of  ore 
extraction  and  in  proportion  to  the  clay  matrix  the  volume  of  ore  may 
be  only  a  fraction.  The  ore  masses  are  rarely  greater  than  200  ft. 
square  horizontally  by  100  ft.  thiek,*but  occasionally  they  are  larger,  as 
in  the  Holbrook  big  stope,  which  is  600  ft.  wide  by  800  ft.  long  on  the 
dip.  The  ore  lenses  are  not  distributed  haphazard  in  the  Escabrosa 
limestone,  but  favor  the  line  of  certain  faults  and  the  porphyry  contact. 

MINING 

Much  of  the  ground  is  soft  enough  to  be  removed  by  a  pick  alone,  but 
it  has  been  found  quicker  to  loosen  it  by  auger  bores  and  blasting.  The 
auger  is  of  the  common  earth  type  with  a  fixed  wooden  handle  and  4  ft. 
long.  The  cutting  diameter  is  the  same  as  that  of  hand  drills,  so  that, 
if  a  boulder  is  struck,  the  hole  can  be  finished  by  single-jacking.  For 
harder  ground  single  jacks  are  usual,  but  experiments  are  now  in  progress 
with  the  hand-hammer  air  drills.  For  hard  rock  3-in.  air  drills  requiring 
two  men  are  in  vogue.  The  blasting  is  always  done  with  35  or  40  per 
cent,  dynamite. 

With  the  gentle  topography  and  flat  ore  horizons  there  are  no  tunnel 
sites,  so  vertical  shafts  are  compulsory.  To  discover  the  ore,  the  ground 
at  the  100-ft.  levels  is  exposed  by  drifts,  following  faults  or  ore  stringers 
where  they  exist,  otherwise  the  area  is  divided  into  more  or  less  rec- 
tangular blocks.  The  volume  of  ground  to  be  searched  is  much  greater 
than  in  vertical  veins,  as  is  shown  by  the  annual  driving  of  around  5  miles 
apiece  by  the  Copper  Queen  and  Calumet  &  Arizona  companies,  and  in 
like  proportion  by  the  others,  besides  considerable  diamond  drilling. 

Drifting. — By  arching  the  roof,  many  of  the  drifts  will  stand  without 
timbering,  but  in  bad  ground  full  sets  with  inclined  posts  8  to  12  in. 
square  and  2  in.  per  foot  batter  are  put  in.  In  the  Calumet  &  Arizona 
mines,  when  the  drifts  penetrate  ore,  their  sets  have  vertical  posts  that 
will  join  with  the  stope  square  sets;  but  in  the  Copper  Queen  the  inclined 
drift  sets  are  kept  in,  until  replaced  by  the  square  sets  of  stoping. 

In  the  Calumet  &  Pittsburg  mine  50  to  60  ft.  advance  in  7  days 
is  made  by  using  a  3-in.  drill  to  get  in  a  round  of  9  to  12  holes,  in  the 


132 


MINING    WITHOUT   TIMBER 


8-hour  shift  from  7  A.  M.  to  3.30  P.  M.,  which  is  then  blasted.  Muckers 
then  come  on  and  begin  to  throw  the  muck  back,  so  as  to  have  the  face 
ready  for  a  new  set  up,  when  the  night  shift  drillers  arrive  at  6  P.  M. 
Tramming  the  mucking  can  then  continue  without  interrupting  the 
drillers  and  when  they  leave  at  2.30  A.  M.,  the  clearing  of  the  face  again 
begins  as  before. 

Sloping. — Except  for  the  Mitchell  system  and  the  stulls  used  in  the 
thin  beds  of  the  Pittsburg  &  Duluth,  the  square-set  system  is  universal 
here  for  timbering  stopes.  For  the  soft,  irregular,  and  patchy  stopes  of 
Bisbee,  the  square-set  system  has  many  advantages.  It  permits  of  the 
easy  omission  of  barren  spots,  gives  the  best  ventilation  and  allows  the 
re-entering  of  an  old  stope  at  any  time  for  the  removal  of  filling,  now 


Fio.  62. — Panel-stoping  at  Copper  Queen  mine. 

valuable  but  not  previously,  or  for  the  -extension  of  the  sets  into  lower- 
grade  ground.  A  junction  with  an  old  stope's  timbers  can  readily  be 
affected  from  any  direction  and  this  is  especially  useful,  when  approach- 
ing from  below,  for  the  timbers  tend  to  prevent  that  sliding  of  filling, 
which  will  occur  with  untimbered  systems.  When  worked  in  small 
panels  and  closely  filled,  square  sets  are  safe  even  in  the  soft  ground  of 
Bisbee. 

The  Copper  Queen  Mines  have  14  posts  between  the  100-ft.  levels. 
The  upper  floor  square  sets  are  framed  7  feet  vertically  by  5  feet  square, 
but  the  sill  floor  sets  are  9  ft.  high  to  allow  space  for  extra  caps  and  for 
settling.  The  stope  floors  are  kept  level  in  spite  of  ground  movement 
by  cutting  the  posts  of  the  exact  length  necessary  to  keep  their  tops 
even  with  that  of  the  corresponding  raise  post.  The  width  of  a  panel 
across  an  ore  body  depends  on  the  hardness  of  ore,  but  varies  from 
5  to  12  sets. 

The  method  of  mining  is  illustrated  by  Fig.  62.  A  cross-cut  c  d 
is  driven  from  the  sill  floor  drift  a  b  and  two-compartment  vertical  raises 


OVERHAND    STOPING    ON    WASTE    IN    THE    UNITED    STATES  133 

c  and  d  are  driven  to  the  top  of  the  ore  body.  One  compartment  is  used 
for  an  ore  chute,  the  other  a  man  and  timberway.  Panel  A  is  taken 
out  floor  by  floor  from  the  top  down,  the  ore  being  thrown  down  the  raise 
to  be  loaded  from  chutes  at  its  base  into  cars  on  the  sill  floor,  the  extra 
waste  for  filling  is  received  from  the  level  above  through  the  raise  before 
mentioned.  The  ore  descending  from  a  floor  is  prevented  from  mixing 
with  the  waste  coming  to  the  floor  from  above  by  a  partition  in  the 
raise,  or  by  the  use  of  an  adjoining  vertical  line  of  square  sets  for  the  ore 
descent.  Only  such  sets  are  kept  unfilled  as  are  necessary  for  the  free 
movement  of  the  miners.  Where  there  are  no  levels  already  existing 
above,  to  which  a  raise  can  be  joined,  a  cage  is  put  in  the  manway  and 
operated  by  an  electric  hoist  set  to  one  side  on  the  level  below.  A  car 
of  waste  can  then  be  raised  on  this  cage  to  the  desired  floor. 

A  number  of  floors  of  panel  A  can  be  worked  simultaneously  in 
benches  by  overhand  stoping;  when  A  is  extracted  and  filled,  panel  B 
can  be  attacked  similarly.  Cross-cut  e  f  and  raises  e  and  /  should  then 
be  ready  for  beginning  panel  C.  About  25  board  feet  of  timber  are  lost 
per  ton  of  ore  extracted;  the  lagging  is  2-in.  plank,  which  is  largely  re- 
covered. The  timber  comes  sawn,  from  the  Pacific  Northwest,  and 
costs  around  $25  per  M. 

Most  of  the  cars,  of  16  cubic  ft.  capacity,  are  of  the  vertical  shaft 
type  having  a  hinge  and  turn  plate  to  enable  them  to  dump  from  one  end 
in  any  direction.  The  Copper  Queen  Co.  employs  simple  turn  sheets 
in  the  drifts;  it  has  installed  electric  traction  and  hoists  most  of  its  out- 
put through  one  shaft. 


CHAPTER  XI 

OVERHAND  STOPING  ON  WASTE  IN  MEXICO  AND 
AUSTRALIA 

EXAMPLE  24. — Los  PILARES  MINE,   NACOZARI,   SONORA,   MEXICO 
(See  also  Example  31.) 

Irregular  Lenses  in  Porphyry.  Two  Methods  of  Sill  Flooring.— 
The  mine  is  at  Porvenir,  five  miles  from  Nacozari,  where  there  is  an 
immense  pear-shaped  " horse"  of  rock  2000  ft.  long  and  having  a 
maximum  width  of  800  ft.  The  horse  a,  as  shown  in  Fig.  65,  is  sur- 
rounded by  ore  deposits  b,  which  vary  in  thickness  from  0  to  200  ft. 
The  interior  of  the  horse  has  been  shattered  and'  at  the  core  there  are 
mineralized  areas  c,  one  of  which  is  300x300  ft. 

The  horse  is  capped  by  a  brecciated,  rhyolitic,  iron-stained  gossan, 
varying  in  thickness  from  20  to  75  ft.  and  carrying  little  or  no  copper. 

Below  the  gossan  is  the  enriched  mineral  zone,  also  of  variable  thick- 
ness, but  averaging  about  100  ft.  In  this  zone,  the  copper  mineral 
changes  from  pseudomorphic  chalcocite  after  pyrite,  to  pyrite  and 
chalcopyrite  with  a  slight  coating  of  chalcocite.  It  is  seldom,  however, 
that  a  complete  replacement  of  pyrite  or  chalcopyrite  by  chalcocite  has 
taken  place. 

Below  the  enriched  zone  are  found  the  primary  sulphides  consisting 
entirely  of  chalcopyrite  and  pyrite.  Throughout  the  entire  ore  body, 
from  the  surface  to  the  lowest  workings,  the  character  of  the  deposit  is 
the  same;  a  shattered  rock  with  the  ores  existing  as  the  cementing  mate- 
rial. Both  the  brecciated  rhyolite  near  the  surface,  and  the  brecciated 
andesite  below  it  contain  copper  minerals. 

There  is  a  dike  x,  y,  z,  Fig.  63,  approximately  bounding  the  south- 
eastern portion  of  the  ore  body,  but  more  and  more  approximately 
traveling  the  center  of  the  ore  deposit  as  it  is  followed  to  the  northwest 
This  dike  of  disintegrated  diabase  has  a  width  varying  in  different  places 
from  a  knife  edge  to  about  30  ft. 

Near  the  surface  this  dike  has  a  slight  dip  to  the  eastward,  making  it 
a  hanging  wall  of  the  ore  body,  but  at  the  300-ft.  level  it  changes  its 
direction  and  dips  to  the  west  with  increasing  flatness  from  about  70  deg. 
on  the  300-ft.  level  to  about  50  deg.  on  the  600-ft.  level.  In  the 
northern  portion  of  the  deposit,  large  spurs  and  splits  from  the  main  dike, 
varying  in  width  from  1  to  25  ft.  and  in  length  from  small  to  great 

134 


OVERHAND    STOPING    ON    WASTE    IN    MEXICO    AND    AUSTRALIA          135 

distances  are  found  running  into  the  eastern  country  rock.  This  dike 
has  caused  great  difficulty  in  the  mining  operations,  owing  to  its  vagaries 
of  direction  and  its  tendency  to  slough  and  cave  from  above. 


FIG.  63. — Plan  of  ore  body,  Los  Pilares  mine. 


From  the  foregoing  it  will  be  understood  that  the  ore  body  has  a 
definite  and  fairly  regular  exterior  boundary.  The  richest  ore  is  apt  to 
lie  very  close  to  this  boundary,  the  grade  of  the  ore  getting  poorer  toward 
the  center  of  the  pear-shaped  horse  until  the  limit  of  commercial  ore  is 


136  MINING    WITHOUT   TIMBER 

reached.  In  what  follows,  the  term  "width"  refers  to  the  distance  from 
the  exterior  boundary  along  a  line  at  right  angles  to  the  same  to  the  point 
at  which  the  commercial  ore  ceases.  This  width  of  ore  varies  in  different 
parts  of  the  mine,  and  on  different  levels,  from  a  few  feet  to  more  than 
200  ft. 

Mining  Methods. — Two  main  working  shafts,  d  and  e,  have  been 
sunk  in  country  rock  about  50  ft.  outside  the  deposit  from  which  it 
has  been  developed.  The  width  and  length  of  the  orebody,  the  kind 
of  ground  found  in  the  deposit,  the  excessive  cost  of  timber  at  the  mine, 
the  cheapness  of  common  labor,  and  several  smaller  items,  were 
the  considerations  which  determined  the  method  of  ore  extraction. 
The  pillar-and-stope  method  is  in  use  throughout  the  workings.  The 
whole  deposit,  or  rather  the  commercial  ore  area,  is  divided  up  into  a 
series  of  stopes  and  pillars,  the  widths  of  which  vary  according  to  the 
width  of  the  ore  and  the  character  of  the  ground  to  be  worked. 

Pillars  and  Pillar  Lines. — The  pillars  are  bounded  by  imaginary 
vertical  planes  extending  from  the  surface  to  the  bottom  of  the  workable 
ore.  Separate  maps  like  Fig.  63  are  kept  up  to  date  for  each  level.  On 
each  of  these  maps,  the  pillars  are  accurately  plotted,  thereby  showing 
the  location  of  every  stope  and  pillar,  its  dimensions,  and  also  the  courses 
of  the  several  pillar  lines. 

When  a  stope  is  to  be  "sill-floored,"  the  engineer  will  set  pillar  plugs 
on  each  side  of  the  stope  that  calls  for  a  pillar.  The  position  of  these 
plugs  will  be  calculated  with  reference  to  their  distances  from  their  cor- 
responding pillar  lines.  The  distance  from  the  pillar  plugs  to  the  pillar 
will  then  be  given  to  the  stope  boss,  and  it  is  his  duty  to  see  that  the  pillar 
line  in  question  is  carried  forward  and  the  plugs  cared  for.  For  a  height 
of  two  or  three  slices,  these  plugs  will  be  changed  by  the  eye  by  the  stope 
boss;  they  will  then  be  checked  by  the  engineer  and  carried  on  as  before. 
These  pillars  vary  in  width  from  25  to  60  ft.  and  are  placed  approximately 
at  right  angles  to  the  country  wall.  Thus,  each  stope  is  bounded  on  two 
sides  by  pillars,  while  the  wall  rock  on  one  end  and  the  end  of  the  com- 
mercial ore  on  the  other  constitute  the  respective  third  and  fourth  sides. 
When  the  ore  zone  is  narrow,  say  from  40  to  100  ft.,  the  above  statement 
applies  and  the  stope  will  vary  from  100  to  150  ft.  in  length  along  the 
length  of  the  ore  zone,  and  pillars  vary  from  25  to  60  ft.  in  length  also, 
measured  along  the  length  of  the  zone,  depending  on  the  nature  of  the 
ground.  When  the  commercial  ore  zone  is  much  wider,  say  250  ft.,  the 
length  of  the  stope  is  either  lessened  and  that  of  the  pillar  increased,  mak- 
ing a  stope  50x250  ft.  and  the  pillar  the  same  size,  or,  in  case  of  excessive 
width  of  ore,  a  third  pillar  is  introduced  running  at  right  angles  to  the 
other  two  pillars  and  practically  making  two  stopes  between  one  set  of 
pillars,  whereas  if  the  ore  had  not  been  so  wide,  only  one  stope  would 
have  been  excavated. 


OVERHAND   STOPING    ON    WASTE   IN   MEXICO    AND    AUSTRALIA         137 

Sill  Flooring. — The  stopes  may  be  "sill-floored"  by  two  methods, 
A  and  B.  By  method  A  the  entire  stope  area  is  cut  out  on  the  level 
floor,  while  by  method  B  a  floor  arch  15  ft.  thick  is  left  above  the  floor 
level,  and  from  the  top  of  this  arch  the  full  stope  area  is  carried  up. 
Method  A  necessitates  either  a  permanent  drift  in  the  pillar  with  cross- 
cuts running  to  the  stopes  and  ending  in  shovelways  or  chutes,  or  neces- 
sitates the  driving  of  permanent  drifts  outside  the  ore  with  cross-cuts 
run  to  chutes  or  shovelways  in  the  stope;  or  both  classes  of  drifts  may  be 
used  for  the  same  stope.  Method  A  is  generally  used  in  bodies  of  high- 
grade  ore,  or  in  weak  ground.  In  method  B  a  permanent  drift  must  be 
maintained  through  the  stope  and  for  that  reason  the  drift  is  protected 
by  the  floor  arch.  This  is  the  method  used  in  the  case  of  wide  and  long 
stopes,  where  the  main  development  drift  has  been  driven  along  the  coun- 
try wall  and  must  be  maintained  in  order  to  extract  the  ore.  With 
method  B  when  the  ore  is  75  or  100  ft.  wide,  an  auxiliary  permanent  level 
is  driven  from  the  main  level  through  the  length  of  the  stope  and  pro- 
tected by  the  floor  arch. 

Three  different  methods  of  stoping  are  in  vogue  in  this  mine.  First, 
square  setting;  second,  overhand  stoping  on  waste  of  this  example,  and 
third,  the  overhand  stoping  with  shrinkage  and  delayed  filling  of 
Example  31. 

Square-set  timbering  and  stoping  is  well  known  and  also  but  little  used 
here,  so  it  will  not  be  taken  up  in  detail.  It  is  adopted  in  soft  ground 
that  is  liable  to  cave.  After  extracting  the  ore  and  timbering  with 
square  sets,  permanent  levels  may  be  either  driven  through  the  pillar  or 
maintained  through  the  center  of  the  stope  by  lagging  over  the  sets 
through  the  center  line  of  the  stope.  The  stope  is  then  filled  with  waste 
up  to  the  top  floor  of  the  square  sets.  Chutes  and  manways  are  carried 
up  by  lining  a  given  timber  set  all  the  way  up  with  3xl2-in.  plank  and 
dividing  it  into  chute  and  manway  compartments  as  in  Example  23. 

-Overhand  stoping  on  waste  may  be  used  after  a  stope  has  been  started 
by  either  of  the  two  sub-methods  (A  and  B)  of  sill  flooring. 

Sub-method  A. — As  an  illustration  of  this  system  with  method  A 
assume  a  stope  a,  Fig.  64,  50  ft.  along  the  orebody  and  200  ft.  wide. 
Corresponding  pillars  6  will  also  be  50  ft.  in  length.  At  50-ft.  intervals, 
cross-cuts  c  will  be  driven  into  the  stope  from  the  pillar  drifts  d  on  both 
sides  of  the  stope  in  question.  Working  from  the  ends  of  these  pillar 
drifts  the  entire  area  of  ore  within  the  pillar  lines  is  removed  by  blast- 
ing to  a  height  of  15  feet.  This  broken  ore  is  trammed  over  tracks 
laid  through  the  several  cross-cuts  from  the  pillar  drifts  and  by  extend- 
ing temporary  tracks  from  them  into  the  center  of  the  stope.  The  bro- 
ken ore  is  shoveled  into  cars  and  taken  out  via  these  tracks.  After  clean- 
ing cut  all  of  the  ore,  the  temporary  tracks  will  be  removed.  Fifteen  feet 
from  the  pillar  b  inside  the  stope,  and  adjacent  to  the  extension  of  the 


138 


MINING    WITHOUT    TIMBER 


pillar,  cross-cuts  c,  are  built  up  the  chute  and  manways  e.  This  15  ft.  of 
cross-cut  is  timbered  after  extracting  the  ore.  The  stope  will  then  be 
filled  with  waste  from  the  mill  holes  /  to  within  a  distance  of  5  ft.  from 


ES  AMD  MINERALS. 


Section  on  A-B 
Fia.  64. — Stoping  by  submethod  "A",  Los  Pilares  mine. 


OVERHAND    STOPING    ON     WASTE    IN    MEXICO    AND    AUSTRALIA          139 

the  roof,  the  work  of  leveling  off  the  waste  from  the  different  mill  holes 
being  done  by  shoveling  and  wheelbarrow  work. 

These  mill  holes/  (see  Fig.  64)  are  adjacent  to  the  ends  of  the  pillar 
cross-cuts.  It  is,  therefore,  obvious  that  they  are  raised  to  the  level 
above  simultaneously  with  blasting  out  the  first  slice.  They  thus  serve 
not  only  for  dumping  down  waste  for  filling,  but  to  ventilate  the  stope. 

Chutes  and  manways  are  kept  built  up  sufficiently  above  the  waste 
packing  so  as  to  prevent  the  filling  from  running  into  them.  The  chutes 
are  generally  cribbed  with  6x6-in.  timbers  with  a  3/4-in.  notch  on  each 
end.  This  leaves  a  4  1/2-in.  opening  in  the  crib  between  each  corre- 
sponding pair  of  timbers.  Large  rock  from  the  packing  is  built  up 
around  the  outside  of  the  chute  to  keep  it  in  place,  and  the  inside  is  lined 
with  3xl2-in.  plank.  The  manway,  usually  about  2  1/2x6  ft.,  is  carried 
up  along  one  side  of  the  chute  and  is  built  out  of  2xl2-in.  plank.  A 
manway  is  sometimes  deemed  unnecessary  for  a  chute,  in  which  case  the 
latter  alone  is  carried  through  the  fill.  In  the  manway,  a  3-in.  diameter  pipe 
is  placed  with  its  top  level  with  the  top  of  the  manway  and  its  bottom 
about  7  or  8  ft.  above  the  floor.  This  serves  the  purpose  of  letting  drill 
steel  down  without  cutting  the  manway  lining  and  breaking  ladders 
which  would  result  from  throwing  down  the  steel.  After  the  filling  has 
been  leveled  off  to  its  proper  height,  another  slice  varying  from  6  to  12  ft.  in 
height  at  the  breast  is  blasted  from  the  stope.  This  slice  is  started  by 
driving  blind  raises  to  the  proper  height  and  enlarging  these  till  they 
intersect,  this  method  enabling  the  drilling  of  flat  water  holes.  It  is  often 
the  case  that  the  same  slice  will  be  started  in  three  or  four  different  parts 
of  the  stope  from  blind  raises.  The  ore  broken  in  this  method  of  slicing  is 
put  into  the  nearest  chutes  by  shoveling  and  wheelbarrow  tramming. 
After  the  slice  has  been  finished,  the  stope  is  cleaned  of  its  ore  and  waste* 
filling  is  run  in  again  and  the  same  procedure  followed  as  before. 

Sub-method  B. — When  the  ore  shoot  is  narrow,  say  30  to  50  ft.  in 
width,  for  quite  a  length  along  the  country  wall,  and  the  ore  is  of  com- 
paratively low  grade,  the  same  method  of  extraction  is  used,  but  with 
the  sill-flooring  of  method  B.  For  this  the  main  drift  is  usually  protected 
by  a  floor  arch,  as  shown  at  i,  Fig.  65.  The  main  drift  is  generally  run 
along  the  country  wall  or  close  to  it  as  at  6.  Cross-cuts  c  are  driven  every 
30  or  40  ft.  at  right  angles  to  the  main  drift  in  the  ore,  the  two  opposite 
term  nal  ones  being  on  the  pillar  lines.  If  the  drift  to  be  maintained  is 
in  the  ore,  then  these  cross-cuts  may  be  driven  at  the  stated  intervals 
on  either  side  of  the  drift,  generally  alternating  from  the  left  to  the  right 
side.  Offset  from  each  cross-cut  and  set  out  8  ft.  from  the  center  of  the 
main-drift  track  a  6  ft.-sq.  raise  is  driven  to  a  height  of  22  ft.  After  raising 
in  each  cross-cut  the  stated  height,  intermediate  drifts  and  cross-cuts 
7  ft.  high  are  run  from  each  raise  making  the  floor  of  each  drift  15  ft. 
above  the  floor  level.  These  intermediate  drifts  are  connected  together 


140 


MINING    WITHOUT   TIMBEK. 


and  the  first  floor  of  the  stope  is  then  excavated  by  enlarging  them  by 
blasting  both  ways  till  they  intersect.  All  the  ore  between  pillar  lines 
is  extracted  to  the  country  rock  wall  in  one  direction  and  to  the  com- 
mercial ore  limit  in  the  other.  Should  the  ore  prove  to  extend  for  20  ft. 
or  more  in  width  from  the  center  of  the  main  drift,  an  8-ft.  pillar  is  left 
between  the  main  drift  and  the  ore,  and  an  auxiliary  drift  d  from  the  cross- 
cuts is  driven  parallel  to  the  main  drift.  From  this  auxiliary  drift  the 
ore  is  sliced  back,  the  broken  material  being  trammed  out  of  the  various 


MINK  ADD  MINERALS 


>  h        h     d     ht 

Section  0/1  J-B" 

FIG.  65. — Stoping  by  submethod  "B",  Los  Pilares  mine. 


cross-cuts  first  driven  from  the  main  haulage  drift.  This  work  leaves  a 
solid  arch  of  rock  8  ft.  wide  on  each  side  and  8  ft.  thick  over  the  main 
haulage  drift  to  protect  it.  When  this  method  is  followed  chutes  are 
cribbed  up  from  convenient  points  about  50  ft.  apart.  For  such  small 
stopes  two  manways  /  usually  suffice,  which  are  placed  adjacent  to  the 
two  chutes  at  opposite  ends  of  the  stope. 

After  the  stope  is  completed  and  cleaned  of  ore,  waste  is  let  into  the 
excavation  and  operations  proceed  as  in  the  previous  case.  If  the  width 
of  the  ore  is  too  narrow  to  admit  leaving  the  ore  pillars  to  protect  the 


OVERHAND    STOPING    ON    WASTE   IN    MEXICO    AND    AUSTRALIA         141 

drift,  it  is,  of  course,  obvious  that  the  first  system  must  be  followed,  after 
which  the  drift  is  timbered  and  waste  filled  in  over  it. 

For  the  filling  of  stopes  one  main  fill  hole  is  raised  to  the  surface  from 
a  level  located  100  to  200  ft.  above  the  stopes  to  be  filled.  This  main 
fill  hole  then  serves  for  from  three  to  five  stopes  or  more.  From  the 
stopes  to  be  filled,  up  to  the  level  on  which  this  main  fill  chute  is  located, 
from  two  to  four  fill  holes  are  driven  for  each  stope.  These  fill  holes  are 
placed  either  beneath  the  center  of  the  track  or  inclined  over  to  the  center 
from  one  side.  Bearer  timbers  are  laid  over  these  holes  and  the  track  run 
over  the  bearers.  A  5-horsepower  electric  motor,  with  a  train  of  10  cars 
of  1  ton  capacity  pulls  the  fill  rock  from  the  main  fill  chute  to  the  particu- 
lar stope  and  fill  hole  where  waste  is  called  for.  In  this  main  fill  hole  at  a 
distance  of  75  to  125  ft.  below  surf  ace  (distance  varying  according  to  the 
point  of  approach  or  entrance)  a  grizzly  station  is  cut  and  a  grizzly  put 
in  over  the  fill  hole.  This  grizzly  is  variable  in  size,  ranging  from  8x20 
ft.  to  10x10  ft.  The  mouth  of  the  fill  hole  is  chambered  outto  a  conve- 
nient size  on  a  side  away  from  the  hole  leading  to  surface.  Bearers 
of  12xl2-in.  timber  spaced  about  3  ft.  apart  are  placed  across  this  hole. 
On  them  are  bolted  40-lb.  rails  so  as  to  form  openings  of  15  in.  square. 
All  rock  falling  on  the  grizzly  must  be  broken  fine  enough  to  pass  through 
these  openings.  When  boulders  come  down  larger  than  15x15  in.,  they 
catch  on  the  grizzly  and  are  broken  up.  This  prevents  the  hole  from 
"hanging  up"  between  this  grizzly  point  and  the  chute  below,  and  also 
eliminates  any  trouble  at  the  chute  in  the  loading  of  the  cars. 

Ore  Transportation. — On  the  main  tunnel  level,  and  located  so  as 
to  reach  all  stopes  advantageously,  six  ore  pockets  or  bins,  similar  to 
Fig.  66,  have  been  cut  out  of  the  rock  within  the  ore  zone,  each  having  a 
capacity  of  from  1,000  to  10,000  tons.  Each  one  of  these  bins  is  pro- 
vided with  from  one  to  three  sets  of  two  chutes  each,  one  set  of 
chutes  filling  a  30-ton  Ingoldsby  bottom-dump  ore  car.  Two  10-ton 
General  Electric  traction  motors  running  in  tandem  pull  a  train  of  from 
six  to  eight  of  these  cars  to  the  main  tunnel  mouth,  Porvenir,  which  is 
the  terminal  of  the  mine  railroad.  From  here  a  60-ton  Baldwin  locomo- 
tive takes  a  train  of  14  cars  to  the  concentrator  at  Nacozari.  On  each 
of  the  succeeding  levels  above  the  main  tunnel  level  as  far  up  as  the 
200-ft.  level,  continuous  connections  have  been  made  to  each  ore  bin  as 
shown  in  Fig.  66.  With  dump  stations  provided  for  each  bin  on  each 
mine  level,  the  ore  from  each  working  finds  its  way  into  the  nearest 
dump.  This  does  away  with  the  hoisting  of  the  ore,  a  costly  item. 

Compensating  Mexican  Labor. — All  work  underground  is  done  by 
contract.  All  development  work  such  as  sinking,  raising,  drifting,  and 
cross-cutting,  is  contracted  to  the  native  Mexicans  at  so  much  per  foot 
driven;  the  company  furnishing  steel,  powder,  fuse,  caps,  etc.,  the 
contractor  only  having  to  keep  his  working  up  to  regulation  size  and  in 


142 


MINING    WITHOUT    TIMBER 


some  cases  running  his  dirt  to  the  chute.  In  case  of  encountering  waste 
in  a  working,  it  will  be  dumped  into  a  fill  stope.  Prices  of  the  work  per 
foot  vary  with  the  kind  of  rock  and  also  depend  on  whether  the  drilling 
is  preformed  by  machine  or  hand.  In  the  stopes  both  machine  and  hand 
drills  are  used,  the  miner  being  paid  so  much  per  foot  drilled  with  a 
machine,  or  so  much  per  foot  with  hand  steel.  In  the  stopes  the  super- 
vision of  the  holes  is  not  limited  to  the  number  drilled  but  they  must  be 
drilled  as  " pointed"  by  the  stope  boss  to  the  stipulated  depth.  Car 
men  receive  so  much  per  car  trammed,  the  price  varying  with  distance 
traveled  and  whether  the  ore  is  shoveled  from  sheet  iron,  or  a  rough 


FIG.  66.— Ore  bins  and  chutes,  Los  Pilares  mine. 


bottom,  or  drawn  from  a  chute.  In  the  stopes,  shovelers  dump  wheel- 
barrows into  chutes  and  are  paid  by  the  number  of  cars  drawn 'from  the 
chutes,  which  are  counted  up  at  the  end  of  the  day.  As  a  check,  the 
height  of  the  ore  in  the  chute  is  taken  before  starting  to  work  and  after 
finishing;  the  number  of  inches  in  the  chute  to  the  car  being  known. 
In  filling  stopes  men  are  given  a  task  of  so  many  wheelbarrow  loads  per 
day  for  a  certain  wage.  All  wheelbarrow  loads  over  or  under  this  num- 
ber are  figured  and  paid  for  in  proportion  of  the  task  to  the  wage.  About 
1,200  men  are  usually  on  the  monthly  pay  roll  with  an  average  daily 
working  force  of  800  men.  From  1500  to  2000  tons  of  ore  is  sent  to  the 
mill  daily  from  the  whole  mine. 


OVERHAND  STOPING  ON  WASTE  IN  MEXICO  AND  AUSTRALIA       143 

EXAMPLE  25. — WEST  AUSTRALIA 

Sub-vertical  Veins  in  Crystalline  Schists;  Rill  Chutes. — When  working 
under  ideal  conditions,  the  ore  body,  if  continuous,  is  divided  into  blocks 
by  equidistant  levels,  about  200  ft.  apart,  and  by  equidistant  winzes  on 
the  hangwall  side  of  the  lode  at  150  ft.  to  200  ft.  apart.  The  reason  for 
having  the  winzes  on  the  hangwall  side  is  that  the  stopes  are.  thus  filled 
with  the  least  handling.  The  diagram  of  Fig.  67  shows  how  a  mine  is 
thus  blocked  out.  It  is  assumed  in  it  that  only  one  winze  is  connected 
with  the  surface  as  a  pass  for  filling,  but  the  number  of  through  passes 
is  purely  a  matter  of  convenience. 

The  filling  usually  consists  of  fresh  residue,  which  may  be  of  sand,  or  of 
roasted  "slimed"  ore  or  of  raw  "slimed"  ore,  and  which  may  contain  up 
to  25  per  cent,  of  moisture.  If  the  residue  contains  too  much  moisture 
there  is  a  danger  of  it  clogging  the  passes,  so  that  sometimes  it  is  necessary 
to  stack  it  on  the  surface  for  a  short  time  previous  to  delivery  to  the 
mine,  and  by  this  means  also  much  of  its  residual  cyanide  content  is 
destroyed.  No  chemical  treatment  whatever  of  the  residue  to  destroy 
its  cyanide  contents  is  practised  in  Western  Australia.  The  residue 
received  from  the  surface  can  be  distributed  to  the  various  winzes  by 
means  of  a  belt  conveyor  system  along  a  disused  level  above  the  stopes. 

The  methods  of  stoping  and  filling  on  the  rill  system  are  as  follows : 

On  any  particular  level  a  leading  stppe  is  taken  out  below  the  ore 
to  be  mined,  when  the  drive  is  timbered  usually,  either  by  single  stulls, 
or  when  the  lode  is  over  14  ft.  wide,  by  saddlebacks,  at  intervals  of  5  ft. 
The  latter  consist  of  pairs  of  stulls  sloping  toward  one  another  like  the 
rafters  of  a  roof,  and  bearing  upon  a  longitudinal  ridging  of  sawn  timber 
2  in.  thick.  The  stulls  are  lagged  with  poles  about  4  in.  in  diameter  of  a 
local  wood  called  gimlet  wood,  or  with  old  iron  pipes.  The  lagging  is  in 
turn  covered  with  old  filter  cloths,  or  the  sides  and  linings  of  cyanide 
cases  or  any  other  inexpensive  material  which  serves  to  prevent  the 
residues  from  falling  through. 

Two  alternative  methods  of  stoping  the  ore  are  shown  at  A  and  B  on 
the  diagram  of  Fig.  67.  In  the  former  all  the  holes  are  "  down"  holes  and 
can  be  drilled  wet,  which  is  an  important  consideration  in  view  of  the 
necessity  of  reducing  dust  production.  In  this  method  the  benches  are 
taken  out  at  an  inclination  slightly  flatter  than  that  of  the  natural  slope 
of  the  filling  which  in  the  case  of  residues  is  about  45  deg. 

In  the  less  common  method  at  B  the  ore  is  mined  by  a  series  of 
horizontal  cuts,  and  .some  of  the  holes  drilled  must  be  "uppers"  and 
drilled  dry. 

Usually  during  the  timbering  of  a  drive  ore  chutes  are  put  in  at  dis- 
tances of  about  50  ft.,  one  (P)  midway  between  the  Winzes  (W),  the 
others  (Q)  being  intermediate.  As  stoping  proceeds  passes  about  4x4 


144 


MINING   WITHOUT   TIMBER 


ft.  in  the  clear  are  built  above  these  chutes,  usually  of  7-in.  logs,  but 
sometimes  of  9x3  in.  sawn  timber.  Each  winze  is  also  " cribbed"  up 
except  when  it  will  not  be  required  later  on  for  passing  "filling"  to  lower 
workings  or  for  ventilation.  In  such  a  case  the  timbering  of  the  level 
below  the  winze  can  be  closed  up  and  the  winze  filled  up. 


Fia.  67. — Stoping  system,  West.  Australia. 


The  breaking  of  the  ore  and  the  filling  of  the  stopes  with  residues 
succeed  one  another  alternately.  Before  the  benches  of  ore  are  blasted 
eucalyptus  saplings  or  slabs  are  laid  on  the  sloping  surface  of  the  residue 
filling.  These  serve  to  keep  separate  to  a  great  extent  the  broken  ore 
from  the  residue,  and  assist  in  its  "rilling"  into  the  passes,  very  little 


OVERHAND    STOPING    ON    WASTE    IN   MEXICO    AND    AUSTRALIA         145 

labor  being  then  required.  The  passes  are  then  built  up  close  to  the 
working  faces  and  covered  over  to  prevent  residues  from  entering  them, 
and  the  poles  or  slabs  are  removed.  More  residue  is  then  dumped  down 
the  winzes  into  the  stopes,  filling  them  up  to  a  convenient  distance  from 
the  faces.  When  a  stope  has  assumed  the  appearance  shown  at  C,  when 
all  ore  can  be  rilled  to  the  passes  P,  the  intermediate  passes  Q  are  no 
longer  entirely  necessary.  As  stoping  proceeds  the  appearance  of  the 
stopes  becomes  similar  to  that  shown  at  D.  It  is  usual  to  stope  a  series 
of  blocks  on  the  same  level  simultaneously,  so  that  the  filling  of  the 
stopes  with  residue  on  both  sides  of  the  winzes  and  the  building  up  of  the 
passes  can  be  carried  on  symmetrically.  When  the  stopes  are  nearly 
beaten  out,  as  shown  at  E,  it  is  usual  to  sink  subsidiary  winzes  R  in  the 
triangular  blocks  of  ore  left  below  the  level  above,  through  which  the 
residues  for  filling  can  be  dropped.  The  lodes  in  Western  Australia  are 
usually  very  steep,  and  this  rill  system  is  generally  particularly  applicable 
to  them.  Generally,  as  the  dip  of  the  lodes  decreases  below  45  deg.,  more 
and  more  shoveling  is  necessary  to  assist  the  rilling,  and  the  method 
becomes  inapplicable  when  the  dip  is  less  than  about  35  deg. 

Filling  with  residue  in  Western  Australia  has  been  in  use  for  about 
13  years,  and  its  present  cost  is  about  lOd.  per  short  ton  of  ore  mined. 

EXAMPLE  26. — BRITISH  AND  OTHER  MINES.     BROKEN  HILL 
DISTRICT,  N.  S.  W. 

(See  also  Examples  27  and  28.) 

Sub-vertical  Vein  in  Crystalline  Schist.  Rill  Chutes.  Cribs  under  a 
Weak  Bach.  Sampson  System. — The  Broken  Hill  district  is  situated  in  a 
desert  country  about  300  miles  west  of  Sydney.  The  great  silver  lead  lode 
occurs  in  the  Barrier  range  which  runs  north  and  south  through  the 
Tertiary  plane  north  of  the  Murray  river  and  its  outcrop  forms  a  narrow 
rocky  line  of  hills  about  11/2  miles  long.  The  main  lode  is  270  ft.  thick 
in  places  (average  about  60  ft.),  and  it  stands  nearly  vertical  and  forks 
in  depth  into  two  slightly  diverging  branches  separated  by  a  horse  of  wall 
rock.  The  present  primary  ore,  now  mined  in  depth,  consists  of  argen- 
tiferous galena  and  zinc  blende  in  a  gangue  of  garnet,  rhodonite  and 
fluorite  of  varying  hardness  and  texture  and  lies  between  walls  of  gar- 
netiferous  gneiss.  The  surface  oxidized  ores  have  been  extracted  mostly 
by  huge  open  pits.  The  debris  from  these  is  now  allowed  to  descend  the 
raises  to  fill  the  stopes  in  the  present  underground  workings.  Mill  tailing 
and  sand  are  also  used  for  filling. 

The  square-set  system,  using  timber  from  Puget  Sound,  was  in  vogue 

for  extracting  the  enriched  oxidized  ores,  which  extended  to  a  depth  of 

300  to  400  ft.,  but  for  the  leaner  sulphides  below  methods  of  stoping 

which  are  cheaper  in  timber  and  less  liable  to  conflagrations  have  been 

10 


146 


MINING    WITHOUT   TIMBER 


introduced.  At  present,  square-setting  is  only  used  in  the  ore  too  friable 
to  stand  by  itself  and  as  an  auxiliary  to  the  systems  of  Examples  25,  26, 
and  27  to  b'e  described.  The  Sampson  system  of  the  British  mine  is 
worked  in  the  following  way  :  After  the  cross-cut  from  the.  shaft  cuts  the 
lode,  it  is  driven  across  to  the  further  wall  and  then  opened  out  on  each 
side  till  the  ore  is  all  taken  out  on  the  sill  floor  from  wall  to  wall  to  a 
height  of  about  11  or  12  ft.  The  face  is  carried  along  in  this  operation 
by  taking  out  the  ore  in  two  stages,  the  bottom  5  or  6  feet  by  drilling 
from  the  sill,  then  rigging  the  machine  on  a  low  bulk,  or  crib,  to  drill  the 
upper  6  feet.  As  the  face  progresses,  the  drive  timbers  of  the  level  are 
put  along  the  footwall  side,  the  leading  set  always  far  enough  from  the 
face  to  be  unaffected  by  the  blasting.  With  the  exception  of  cribs  here 


FIG.  68. — Cribbing  back  of  stope,  British  mine. 

and  there  the  back  will  stand  unsupported.  The  cribs,  built  up  from  the 
floor  and  also  from  the  drive  timbers,  as  shown  in  Fig.  68,  are  built  up 
with  lOxlO-m.  Oregon  timbers,  6  ft.  or  7  ft.  long.  When  within  a  foot 
or  so  of  the  back,  long  stringers  of  10x6  in.  or  10x10  in.  are  cantilevered 
out  (if  necessary  on  every  side)  to  take  in  any  bad  ground  near  top  of  the 
crib,  the  top  is  blocked  and  wedged  tightly  against  the  rock  back,  and 
the  whole  crib  finally  tightened  by  driving  in  wedges  between  its  lower 
timbers,  with  a  spalling  hammer. 

The  drive  timbers  are  shown  at  the  right  in  Fig.  68,  and  by  the 
drawing  Fig.  69.  The  sills  and  struts  are  15  ft.  long,  taking  three  sets 
spaced  on  5-foot  centers.  The  sills  butt  midway  between  the  sets,  while 


OVERHAND    STOPING    ON    WASTE   IN   MEXICO    AND    AUSTRALIA         147 

the  struts  meet  over  a  leg.  The  leg  is  tennoned  into  the  sill  and  strut, 
and  below  the  sill  at  each  leg  a  good  block  is  put  into  the  solid  bottom. 
The  cap  is  jointed  into  the  struts  and  legs.  Along  each  strut  a  15-ft. 
lOxlO-in.  stringer  is  laid,  thus  giving  20x10  in.  of  strut  timber;  these 
stringers  being  properly  spread  by  a  10x2  in.  spiked  to  the  top  of  the 
cap  while  10x6-in.  top  laths,  laid  close  together  from  stringer  to  stringer 
as  shown,  give  support  to  the  filling  placed  on  them.  It  will  be  seen  that 
there  is  an  8-in.  space  below  the  top  laths,  this  being  intended  to  prevent 
the  weight  of  the  filling  from  coming  on  the  cap,  throwing  it  instead  on 
to  the  struts  and  legs.  The  weight  is  thus  kept  to  the  sides  of  the  drive, 
the  cap  merely  taking  part  of  the  side  pressure.  On  the  outside  of  the 
legs  a  vertical  10x4  in.  takes  the  10xl2-in.  horizontal  lagging,  15  ft.  in 
length.  The  reason  for  this  10x4  in.  is,  that  when  the  stope  is  high  the 
weight  of  the  filling  becomes  so  great  that  it  is  necessary  to  put  in  extra 
legs  to  support  the  strut  between  the  sets,  and  as  the  side  pressure  causes 
the  lagging  to  bulge  inward,  the  10x4  will  allow  it  4  in.  play  (this 
has  been  found  no  more  than  sufficient)  before  the  lagging  trespasses  the 
plane  of  the  back  of  the  legs.  Thus,  extra  legs  can  be  put  in  without 
cutting  away  the  bulged-in  lagging  or  coming  within  the  drive  space. 
The  lower  10x4-in.  spreader  butts  2  in.  on  the  leg.  and  2  in.  on  the  sill, 
is  well  packed  underneath,  and  its  top  side  forms,  with  10x2-in.  pieces  laid 
from  sill  to  sill,  the  floor  on  which  the  rails  are  laid.  Outside  the  sets, 
over  the  whole  of  the  sill  floor  are  laid,  parallel  to  drive,  10x4-in.  pieces, 
15  or  16  ft.  long,  butting  against  one  another,  on  which  the  filling  is  laid. 
It  will  thus  be  seen  that  when  coming  up  with  a  stope  from  below,  the 
miners  will  have  no  timbers  shorter  than  15  ft.  long  to  catch  up,  when 
the  last  bit  of  ground  is  being  taken  out. 

At  intervals  of  from  80  to  100  ft.  cross-cuts  are  put  out  from  the  main 
drives  along  the  foot  to  the  hanging  wall.  At  30-ft.  intervals  along  the 
lode  drives,  the  two-compartment  timber  sets,  Fig.  69,  are  put  in  butting 
on  the  drives.  A  set  forms  the  bottom  for  the  ladderway  and  ore  chute, 
built  up  from  the  sill  as  the  height  of  the  stope  increases,  the  top  set  being 
kept  level  with  the  filling.  The  sets  of  lOxlO-in.  pieces  are  laid  one  upon 
the  other  and  held  together  by  the  filling.  When  a  stope  has  been  carried 
up  about  50  ft.,  the  ore  chute  is  by  then  very  much  worn,  and  so  the 
ladderway  and  chute  are  interchanged. 

When  all  these  timbers  are  in  on  the  sill  floor,  the  filling  is  run  in  all 
around  them  and  the  stope  filled  to  within  a  few  feet  of  the  rock  back, 
then  starting  from  the  winzes  (sunk  about  100  ft.  apart),  another  hori- 
zontal strip  is  taken  off  in  the  same  way  for  10  or  12  ft.  in  height,  the 
back  being  supported  where  necessary  with  cribs.  The  waste  filling  sent 
down  the  winzes  from  the  level  above  is  built  out  around  the  timbering 
by  trucking  from  the  winze  chutes  in  cars  running  on  temporary  tracks 
extending 'from  the  winzes.  When  the  filling  comes  to  a  crib,  another 


148 


MINING    WITHOUT   TIMBER 


low  crib  is  built  on  the  new  filling  as  near  as  possible  to  the  old  one  and  the 
latter  then  pulled  down,  and  the  crib  timbers  are  used  over  and  over 
again.  The  chutes  and  ladderways  are  built  up  level  with  the  filling  as 
the  stope  progresses  as  before  stated.  In  this  way  the  ore  is  taken  out 
from  wall  to  wall  till  the  back  is  60  ft.  above  the  sill. 

It  is  then  not  considered  safe  to  carry  the  back  further  up  by  horizontal 
stripping,  but  instead,  starting  from  the  winzes,  the  ore  is  taken  out  in 
diagonal  strips  or  " rills,"  illustrated  by  Fig.  70,  the  slope  of  the  back 
being  approximately  that  of  the  filling 's  slope  of  rest.  By  this  method 
of  stoping  the  only  weak  place  is  the  part  where  the  top  of  the  stope 
meets  the  level  above.  This  point  has  to  be  well  supported  by  an  extra 
number  of  cribs,  also  the  points  over  the  edge  of  the  filling  which  will  by 


SO"  A  '6"  TopLal-hs 


IO"A  2"  Spreader 


"X  4"  Spreade 


T 


<5///  Piece 
FIG.  69. — Drift  timbering,  British  mine. 

this  have  reached  within  6  or  7  ft.  of  the  level.  The  old  10x4-in  bottom 
timbers  are  caught  up  by  a  line  of  cribs  spaced  on  about  10-ft.  centers 
extending  from  wall  to  wall,  the  bottom  timbers  being  directly  supported 
by  long  lOxlO-in.  stringers  laid  across  the  tops  of  these  cribs.  In  this 
way  the  old  bottoms  are  all  supported  by  the  cribs,  which  are  left  in, 
and  the  filling  shoveled  up  around  them  as  close  as  possible.  Thus,  the 
ore  is  all  taken  out,  even  the  last  pyramid-shaped  piece  of  ground  prac- 
tically unsupported  by  the  walls  not  presenting  any  special  difficulty. 
In  the  British  mine  this  method  has  been  very  successful.  The  lode 
there  does  not  attain  as  great  a  width  as  in  some  of  the  other  mines,  its 
greatest  width  being  130  ft.,  while  the  average  width  is  not  more  than 
70  or  80  ft.  The  walls  are  both  firm  and  the  ore  throughout  of  a  consist- 


OVERHAND   STOPING    ON    WASTE   IN   MEXICO    AND    AUSTRALIA         149 

ently  compact  character,  so  that  no  other  method  has  been  required. 
The  filling  used  which  was  formerly  mill  tailings  is  now  zinc  plant  residues. 
Placed  in  a  damp  condition  and  having  greater  mobility  than  waste,  it 
sets  better  and  finds  its  way  into  the  far  corners,  so  that  shrinkage  is 
reduced  to  a  minimum. '  On  these  accounts  the  British  mine  has  mined 
cheaper  than  any  of  the  others,  the  underground  costs  for  some  years 
going  about  9  shillings  per  ton. 

In  all  the  other  mines  open  stoping  with  temporary  cribs  has  been 
adopted,  but  the  other  details  usually  have  been  modified,  In  some 
cases,  struts  in  the  drive  sets  are  short,  only  reaching  from  leg  to  leg,  and 
the  sills  are  put  across  the  drive  instead  of  parallel  with  it.  In  the  Block 
10  mine,  when  coming  up  under  old  bottoms,  the  whole  of  the  last  40  ft. 
is  taken  out  by  square  sets  started  off  the  filling,  the  old  bottoms  on  level 
above  being  caught  up  from  the  top  sets. 


MINES  AND  MINERALS 

FIG.  70. — Longitudinal  section  of  stope,  British  mine. 

In  the  Proprietary  mine  there  are  several  big  differences  in  the  work- 
ing of  this  system,  some  of  them  due  to  the  greater  width  of  the  lode 
worked.  In  the  instance  described  here,  the  lode  has  a  width  of  200  ft. 
In  the  first  place  a  drive  is  put  in  the  footwall  alongside  the  lode  about 
20  or  30  ft.  in  the  country  rock  and  cross-cuts  at  20-ft.  intervals  off  this 
drive,  as  shown  in  Fig.  70.  The  sill  floor  is  taken  out  in  the  manner 
already  described,  and  eventually  these  cross-cuts  tap  the  main  timbered 
drive  which  runs  down  the  middle  of  the  lode.  The  drive  in  the  country 
is  an  advantage  in  case  the  timbered  drive  collapses  at  any  point. 

The  chutes  and  ladderways,  formed  by  a  double  square  set,  are  placed 
about  30  ft.  apart  along  the  drives.  The  double  set  is  close  lagged  all 
around  outside  with  horizontal  10x2-in.  pieces,  and  the  inside  of  the  ore- 
chute  set  is  lined  with  vertical  10x4-in.  hard  wood  (stringy  bark)  butting 
and  spiked  at  the  caps  and  struts  of  the  sets.  This  is  the  only  case  where 
other  than  Oregon  timber  is  used  below  ground.  The  hardwood  makes 


150 


MINING    WITHOUT    TIMBER 


an  excellent  chute  lining,  as  the  wear  is  very  slight  and  the  polished  sur- 
face soon  acquired  is  of  benefit,  whereas  the  Oregon  timber  is  soon  cut 
away  and  the  chips  give  trouble  in  the  concentrating  mill.  If  one  of  the 
lining  pieces  is  knocked  away,  the  outside  10x2-in.  lagging  prevents 
rilling  from  running  into  chute.  It  should  be  mentioned  that  during  the 
last  few  years  the  Proprietary  has  used  for  the  major  part  of  its  mine 
filling  the  sand  residues  from  the  zinc  plant.  These  necessitate  close 
lagging;  where  spaced  lagging  is  mentioned,  the  filling  is  waste. 

On  account  of  the  width  of  the  lode,  part  of  it,  in  section  somewhat 
less  than  half  the  width  as  shown  in  Fig.  71  is  left  as  a  solid  pillar  except 
where  cross-cuts,  etc.,  cut  it  on  the  sill  floors  and  only  the  lode  on  the 


FIG.  71. — Cross-section  of  stope,  Proprietary  mine. 

footwall  side  is  worked  out  for  the  time  being.  A  line  of  chutes  forms 
the  boundary  of  the  stope  at  the  pillar.  The  stope  is  taken  up  in  the 
usual  way  by  cribs,  etc.,  till  the  back  is  75  ft.  above  the  sill.  When  the 
stope  is  being  filled,  vertical  10x4-in.  timbers  are  placed  6  ft.  apart  against 
the  pillar  and  across  these  as  the  filling  rises  are  placed  5x2-in.  "  paddock 
laths"  a  few  inches  apart,  the  intervening  spaces  being  covered  with 
waste  pieces  of  candle  boxes  and  saw-mill  "flitches." 

The  remaining  25  ft.,  up  to  the  level  above,  is  removed  in  the  manner 
illustrated  by  Fig.  72. 

At  each  of  the  waste  winzes,  which  are  about  every  100  ft.  apart, 
a  cross-cut,  16  ft.  wide  by  8  ft.  high,  untimbered  except  for  cribs,  is 
taken  out  across  the  block  from  wall  to  pillar  and  is  then  tightly  filled 
from  the  winzes.  On  this  filling  another  cross-cut  is  driven  about  8  ft. 
high  and  7  ft.  wide  and  also  filled,  the  sides  being  first  chamfered  down 
to  the  sides  of  the  16x8-ft  cross-cut.  Above  this  cross-cut  another  one 
8x7ft.  is  driven,  this  time  timbered  with  "  clap-me-down  sets,"  which 


OVERHAND    STOPING    ON    WASTE    IN    MEXICO    AND    AUSTRALIA 


151 


catch  up  and  support  on  their  tops  the  bottom  of  the  level  above  as 
shown  in  Fig.  72.  These  sets  are  now  filled  up  as  close  to  the  under  side 
of  level  as  possible,  only  the  set  alongside  the  pillar  being  always  left 
unfilled.  In  filling  the  sets,  a  10x2-in.  board  is  put  along  each  side  of  the 
drive  against  the  bottom  of  the  legs,  and  the  struts  take  the  ends  of 
vertical  paddock  laths,  as  before  explained,  the  intervening  spaces 
between  filled  as  before  with  candle-box  pieces,  which  serve  to  hold  the 
filling.  Stoping  across  the  block  now  ceases,  but  is  resumed  in  the 
direction  of  the  lode's  strike,  by  blasting  down  with  a  sloping  breast 
(Fig.  72)  from  the  side  of  this  top  run  of  sets  to  the  under  side  of  the 
25-ft.  deep  block.  The  breasts,  which  extend  right  across  the  block 


r/evel) 


FIG.  72. — Long,  section  of  stope,  Proprietary  mine. 

and  are  worked  from  chute  to  wall,  are  carried  along  from  one  cross- 
stope  to  another  till  the  block  is  worked  out.  As  in  the  case  at  the 
British  mine,  the  top  of  the  stope  is  the  weak  part,  and  has  to  be  well 
supported  by  the  sets,  cribs  also  being  put  in  under  the  back  whenever 
necessary.  When  the  breast  has  advanced  sufficiently,  another  cross- 
run  of  timber  sets  is  put  in  alongside  the  first.  The  filling  for  the  first 
sets  now  has  to  be  dumped  down  where  the  drive  above  crosses  the  top  of 
the  stope.  The  set  next  the  pillar,  as  before,  is  always  left  open,  so 
that  when  the  footwall  block  is  worked  out,  there  is  just  below  the  level 
alongside  the  pillar  an  open  drive  of  sets  which  connects  all  the  middle 
ladderways  and  chutes.  By  this  means,  openings  are  left  for  attacking 
the  pillar  which  will  probably  be  removed  in  a  somewhat  similar  manner, 
depending  on  conditions  existing  when  the  time  (not  yet  arrived)  comes 
for  working  them.  If,  at  any  place,  the  breast  is  too  broken  to  be  worked 
safely  by  the  open  method,  the  lower  face  of  the  block  is  caught  up  on 
cross-cuts  timbered  with  "  clap-me-down  sets,"  stepping  them  up  and 
backward  to  the  top  run  of  sets. 

In  some  cases,  instead  of  working  out  the  ore  to  a  certain  height  by 
horizontal  slices,  it  is  worked  out  by  the  sloping  breasts  entirely.  This 
method  has  two  great  advantages  in  that  nearly  all  the  drill  holes  re- 
quired are  down  holes  and  the  waste  being  on  the  slope,  the  filling  is 
easily  and  cheaply  done.  Where  the  ore  lies  in  flat  layers,  the  back  will 


152 


MINING    WITHOUT   TIMBER 


probably  support  itself  better  when  worked  by  slopes  than  when  it  is 
mined  in  horizontal  slices.  On  the  other  hand,  it  is  more  difficult  to 
support  the  sloping  back  by  cribs.  A  place  has  first  to  be  prepared  on 
the  waste  sill  for  laying  the  first  timbers,  and  the  sloping  back  does  not 
come  squarely  on  the  top  of  the  crib.  The  waste  probably  does  not  set 
so  tightly  on  the  sill  as  in  other  cases  where  much  trucking  has  to  be  done, 
but  still  where  the  walls  are  firm  and  the  back  is  good  and  does  not  re- 
quire much  support,  this  method  can  be  advantageously  adopted.  It 
has  been  used  extensively  in  the  South  Mine  where  conditions  are  favor 
able.  Here  an  extra  number  of  ore  chutes  are  put  in,  and  the  cribs  are 
often  built  up  from  the  tops  of  some  of  these. 

EXAMPLE  27. — PROPRIETARY  MINE,  BROKEN  HILL,  N.  S.  W. 
(See  also  Examples  26  and  28.) 

Sub-vertical  Broken  Vein  in  Crystalline  Schist.  Cross-cutting  in 
Panels. — In  the  Proprietary  Mine  a  class  of  ground  was  met  with  in 
which  the  hard  sulphide  ore  was  found  to  be  broken  up  into  big  and  little 
boulders.  It  was  impossible  to  work  this  by  the  open  method,  and  if 
square  sets  were  used,  the  sudden  loosening  of  a  big  boulder  would  be 
likely  to  knock  over  a  dozen  or  more  sets,  with  disastrous  results.  A 
system  was  therefore  adopted  in  which  as  small  an  area  of  the  back  as 

possible  was  left  unsupported  while  work- 
ing out  the  ore.  The  piece  of  ground 
worked  in  this  way  was  approximately 
60x60  ft.  To  illustrate  this  method  we 
will  consider  that  one  floor  shown  in  Fig. 
73,  is  just  finished,  A,  B,  and  C  are  the 
jump-offs  of  ladderway  and  chute,  while  M 
is  the  waste  winze.  A  is  first  raised  one 
floor  and  a  drive  started  off  toward  B  con- 
necting on  its  way  with  C.  When  drive 
A  B  is  complete  a  cross-cut  is  put  out  to  the 
waste  chute.  Two  men  are  then  started  at 

say  a  to  cross-cut  toward  the  wall,  and  when  about  a  third  of  the  way  in, 
another  pair  start  at  say  b.  When  a  is  two-thirds  of  the  way  in,  a  third 
party  starts  at  c.  When  the  cross-cut  a  reaches  the  wall,  the  miners  are 
withdrawn  to  start  another  cross-cut  and  the  filling  gang  is  put  in.  This 
consists  of  two  truckers  and  one  shoveler,  the  three  working  on  contract, 
together  with  one  man  on  wages  to  look  after  the  packing.  The  cross- 
cut is  filled  in  a  manner  similar  to  that  previously  described.  The  filling 
is  shoveled  up  as  tightly  against  the  rock  back  as  possible,  and  by  the 
time  this  cross-cut  a  is  filled,  cross-cut  b  will  be  ready  and  so  on.  Toward 
the  end  the  corss-cuts  will  be  put  in  alongside  the  filled  ones.  Usually 


FIG.  73. — Plan  of  Gross-cut  slope, 
Proprietary  mine. 


OVERHAND   STOPING    ON    WASTE    IN   MEXICO    AND    AUSTRALIA         153 

the  cross-cuts  on  the  A  side  of  the  waste  chute  are  finished  before  those  on 
the  other  side,  and  while  those  on  the  B  side  are  being  finished,  the  drive 
on  the  A  side  is  filled  up,  the  jump-off  A  is  raised  a  set  and  the  drive  again 
started  on  floor  above  toward  B  as  before.  When  the  cross-cuts  on  the 
B  side  are  all  filled,  the  drive  is  filled  and  lastly  the  waste  cross-cut  to 
M  is  filled  and  thus  a  floor  about  8  feet  thick  is  completely  taken  out. 

The  drive  sets  used  in  this  method  are  shown  in  Fig.  74.     The  legs  of 
the  set  are  about  7  feet  long,  are  tapered,  being  10x6  in.  at  top,  and  8x4 


FIG.  74. — Timbering  of  cross-cut  stope,  Proprietary  mine. 

in.  at  bottom  and  have  a  hole  of  1-in.  diameter  about  6  in.  from  the  top. 
These  legs  foot  on  to  a  10x4  in.  scrap  piece  about  18  in.  long.  The  cap  is 
10x10  in.,  its  ends  simply  laid  on  the  tops  of  the  legs,  while  across  the 
caps  is  laid  whatever  timber  is  necessary  to  support  the  rock  back  (this 
at  times  is  a  fairly  high  crib).  A  10x2  in.  spreader  butting  half  on  the 
cap  and  leg  is  held  up  by  rough  cleats  nailed  to  the  top  of  the  legs.  When 
the  cross-cut  timbers  are  put  in,  the  drive  caps  are  extended  along  the 
sides  of  the  cross-cut,  becoming  struts,  as  shown  in  Fig.  74  (6),  and  butt 
over  a  10x4  in.  corbel  on  top  of  the  leg.  Across  the  struts,  10x10  in.  caps 
are  laid  with  timbers  on  these  to  catch  up  the  back.  No  spreader  is  used 


154  MINING    WITHOUT    TIMBER 

in  the  cross-cut,  and  where  the  pressure  at  end  of  one  is  considerable, 
diagonals  are  put  in  between  sets  as  shown.  The  sets  are  approximately 
6  ft.  wide  and  6  ft.  apart,  but  there  dimensions  are  not  strictly  adhered 
to,  as  the  sets  are  only  of  a  temporary  character.  If  a  boulder  bulges  into 
the  drive  the  set  may  be  made  narrower  to  suit,  or  if  available  timbers 
are  longer  or  shorter  than  necessary,  the  sets  are  built  to  suit,  the  size  of 
the  sets  being  always  subordinate  to  the  ground  and  timber.  When 
opening  out  a  new  floor,  some  of  the  timber  is  recovered  from  the  floor 
below  usually  all  the  caps  and  some  of  the  legs  and  other  parts  being 
saved,  the  hole  in  the  tapered  leg  being  used  to  put  in  a  drill  or  bar  to 
assist  in  withdrawing.  About  two-thirds  of  all  the  timber  used  is  re- 
covered. In  this  system  it  is  desirable  to  drill  by  hand  labor,  using 
shallow  holes  and  light  charges,  on  account  of  the  unstability  of  the  ground. 
By  the  above  means,  this  difficult  ground  is  not  only  safely  worked,  but 
at  a  cost  little  more  than  for  ordinary  ground  with  square-setting. 


CHAPTER  XII 

OVERHAND   STOPING  WITH  SHRINKAGE  AND 
DELAYED  FILLING 

EXAMPLE  28. — CENTRAL  MINE,  BROKEN  HILL,  N.  S.  W. 
(See  also  Examples  26  and  27.) 

Sub-vertical  Veins  in  Crystalline  Schists. — Auxiliary  Cribbing  and 
Square-setting. — In  this  mine  the  lode  reaches  its  greatest  and  most 
constant  width,  and  it  was  not  considered  feasible  to  work  the  immense 
orebody,  2  per  cent,  solid  ore,  by  the  ordinary  systems  in  vogue  at 
Broken  Hill.  The  method  adopted,  as  illustrated  in  Fig.  75,  was  to 
divide  the  lode  up  into  50-ft.  sections  by  vertical  planes  running  across 
the  lode  and  working  out  every  alternate  section  as  a  stope,  leaving  the 
others  as  pillars  till  the  stopes  are  worked  out  and  filled.  These  tilled 
stopes  act  then  as  pillars  while  the  former  pillars  are  worked  out.  This 
is  carried  out  in  the  following  way :  A  main  drive,  timbered  by  8x6x6ft. 
sets  is  first  driven  approximately  down  the  center  of  the  orebody  to 
certain  section  lines  and  this  drive  afterward  connected  by  cross-cuts 
to  a  waste  drive  in  the  footwall. 

At  every  stope  block  the  whole  of  the  sill  floor  is  taken  out,  using 
square  sets  from  wall  to  wall.  A  winze  is  sunk  from  level  above  about 
in  the  middle  of  the  side  of  each  stope,  half  the  winze  in  the  pillar  and 
half  in  stope,  as  shown  at  W,  Fig.  75.  The  cross-cut,  or  gangway, 
timber  sets  next  the  pillars  are  left  open  and  the  row  of  sets  joining  the 
ends  of  these  are  also  left  unfilled.  All  the  other  inside  sets  are  filled 
with  waste  excepting  the  chute  sets  which  are  started  in  the  rows  next 
the  gangways.  The  stope  is  then  taken  up  from  the  top  of  these  sets  by 
the  overhand  stope  and  crib  method  of  Example  26  with  the  difference 
that  the  gangway  row  of  sets  is  carried  up  on  each  side  of  the  stope  and 
left  unfilled,  acting  as  a  barricade  to  keep  the  waste  clear  of  the  pillar, 
and  also  as  a  place  from  which  the  pillar  may  be  attacked.  The  rock 
pack  of  the  stope  is  given  a  slight  slope  downward  from  the  winze  side 
to  the  opposite  side.  When  the  back  has  a  height  of  60  or  70  ft.  above 
the  sill,  square  sets  are  started  on  the  waste,  and  the  rest  of  the  stope 
is  taken  out  in  this  way.  So  great  are  the  ore  resources  of  this  mine 
that  only  a  few  of  the  stopes  have  been  worked  out  and  no  necessity  has 
yet  arisen  for  working  the  pillars  on  a  large  scale. 

The  intended  method  for  working  them  out,  however,  is  indicated 

155 


156 


MINING    WITHOUT    TIMBER 


in  Fig.  75,  at  M.  Starting  from  the  hanging  wall,  the  part  furthest  from 
the  country  drives  and  the  shaft,  a  row  of  sets  is  put  in  against  the  wall 
from  one  filled  stope  to  next,  and  this  row  is  carried  up  from  one  level 
to  the  one  above.  This  is  done  from  every  level  and  at  every  pillar 
simultaneously  as  nearly  as  possible.  These  sets  are  all  filled  except 
the  last  set  of  the  row  on  the  winze  side.  This  one  is  made  the  waste 
chute  from  which  to  fill  the  next  row  taken  out.  In  this  way  the  pillars 
are  gradually  sliced  away  from  top  to  bottom  by  vertical  strips  parallel 
to  the  drives,  working  from  hanging  to  footwall  till  only  worked-out 
ground  is  left  behind.  Enough  work  has  been  done  in  this  way  to 


MIKES  »»C  MINIBUS 


^gWa// — - 

FIG.  75. — Plan  of  sloping,  Central  mine. 

demonstrate  that  under  ordinary  circumstances  the  pillars  can  be  suc- 
cessfully worked  in  this  way,  but  on  account  of  the  heavy  moving  ground 
characteristic  of  the  lode  at  this  part  of  it,  it  is  probable  that  the  system 
when  worked  on  a  large  scale  will  have  to  be  modified.  The  room-cavieg 
system  of  Example  46  seems  to  the  author  more  applicable  for  mining 
these  pillars,  as  slicing  horizontally  could  be  more  easily  controlled 
than  slicing  parallel  to  a  steep  and  heavy  hangwall. 

EXAMPLE  29. — KING  MINE,  GRAHAM  COUNTY,  ARIZ. 
(See  also  Examples  22,  30  and  40.) 

Irregular  Lenses  in  Porphyry;  Auxiliary  Back-caving  and  Underhand 
Sloping. — The  two  lenticular  oreshoots  of  the  mine  are  700  and  500  ft. 
long  respectively  and  fill  a  fault-fissure  in  a  granite  porphyry  hill.  The 


OVEKHAND    STOPING    WITH   SHRINKAGE    AND    DELAYED    FILLING       157 

faulting  has  been  severe,  but  in  the  absence  of  any  sedimentary  rocks,  the 
amount  of  displacement  cannot  be  determined.  The  ore  is  chalcocite 
and  chalcopyrite  in  a  gangue  of  brecciated  granite  porphyry  and  varies 
in  width  up  to  30  ft.  The  vein  dips  at  an  angle  of  70  deg.  and  the  walls 
are  strong  and  well  defined.  The  steep  slope  of  the  mountain  permits 
of  the  vein  being  worked  from  adit  levels,  the  lowest  of  which  gives 
a  vertical  depth  of  600  ft.  below  the  outcrop. 

HAULAGE  ROADS 

Main  haulage  roads  are  driven  in  the  foot-  and  hanging-walls,  parallel 
with  the  vein,  but  at  a  distance  of  from  15  to  20  ft.  from  it.  From  these 
roads,  cross-cuts  are  made  at  intervals  of  25  ft.,  those  in  the  hanging-wall 
being  staggered  or  spaced  midway  between  those  in  the  foot-wall  as  in 
Fig.  76. 


Plan 


Section. 

The  Engineering  $  Mining  Journal 


FIG.  76. — Sloping  at  King  mine. 


The  ore  is  then  broken  from  wall  to  wall  for  the  whole  length  of  the 
ore-shoot.  The  broken  ore  is  at  first  shoveled  out,  but  as  stoping  pro- 
gresses it  is  allowed  to  accumulate,  sufficient  being  removed  to  allow  a 
working  space  of  6  ft.  between  the  broken  ore  and  the  roof.  Two-thirds 
of  the  broken  ore  is  left  till  the  stope  is  worked  out,  the  ore  serving  as  a 
working  floor  for  the  miners  and  also  prevents  caving  of  the  walls. 

OVERHAND  STOPING 

Access  to  the  stope  is  obtained  from  raises  made  in  the  roof  at  inter- 
vals of  100  ft.  and  connected  to  an  upper  level.  From  these  raises,  the 
roof  is  broken  in  horizontal  slices  of  from  10  to  15  ft.  in  thickness.  As 


158  MINING    WITHOUT    TIMBER 

the  miners  work  outward  from  the  raises,  the  sag  or  belly  of  ore  between 
generally  breaks  off,  leaving  the  roof  sufficiently  arched  to  allow  the 
block  to  be  broken  from  on  top.  Large  horses  of  hard,  barren  ground 
frequently  occur  in  the  vein  and  these  are  left  in  as  pillars,  to  support 
the  walls. 

Occasionally,  parts  of  the  vein  are  too  soft  to  be  mined  safely  by 
overhand  stoping  and  the  mode  of  attack  is  changed.  From  the  two 
raises  between  which  the  ore  is  softer  than  usual,  a  drift  is  made  20  to 
30  ft.  above  the  back  of  the  stope  and  connecting  the  raises.  Midway  in 
this  drift  down-holes  are  drilled  in  the  floor  and  sides.  As  these  holes 
are  blasted  and  break  down  the  shell  of  ore  between  the  floor  of  the  drift 
and  the  stope,  mining  is  continued  back  to  the  raise  until  the  whole  of  the 
shell  has  thus  been  broken  by  underhand  stoping.  In  using  this 
method,  the  roof  of  the  drift  which  connects  the  raises  must  be  sufficiently 
high  to  allow  the  handling  of  the  long  jumper  drills  needed  in  breaking 
down  the  floor.  When  approaching  an  upper  level,  the  ore  is  always 
broken  by  underhand  stoping. 

When  the  top  of  the  orebody  is  reached  in  stoping,  the  remainder  of 
the  broken  ore  is  drawn  off  through  the  cross-cuts.  A  certain  admixture 
of  wall  rock  and  ore  is  unavoidable  when  the  last  portion  of  the  ore  is 
drawn,  but  this  is  easily  removed  on  the  sorting -platform  over  which  the 
ore  is  passed.  The  switches  at  each  cross-cut  on  the  road-way  allow  the 
shovelers  to  load  their  cars  without  interfering  with  the  haulage.  The 
ore  is  sledged  and  loaded  by  contract,  and  when  a  car  is  full  it  is  pushed 
into  a  side  track,  where  the  mule  train  is  made  up.  A  few  miners  are 
required  to  block-hole  the  larger  pieces  of  ore  as  they  appear  at  the 
shoveling  openings.  The  advantages  of  this  method  are  obvious;  one 
worthy  of  special  notice  is  the  security  in  which  the  shoveler  works. 

EXAMPLE  30. — CORONADO  MINE,  GRAHAM  COUNTY,  ARIZ. 
(See  also  Examples  22,  29  and  40.) 

Irregular  Lenses  in  Porphyry;  Auxiliary  Back-caving  of  Rooms,  and 
Subsequent  Pillar-caving. — This  mine,  one  of  the  most  important  holdings 
of  the  Arizona  Copper  Company,  lies  on  the  southern  slope  of  theCoronado 
mountain,  a  granite  massif,  whose  precipitous  sides  form  a  conspicuous 
landmark  in  the  district.  The  great  Coronado  fault  is  at  the  base  of  these 
granite  bluffs.  It  strikes  east  and  west  and  can  be  traced  for  a  length  of 
two  miles.  Its  movement  has  been  downward  and  \\esterly  at  its  eastern 
extremity,  resulting  in  a  vertical  displacement  of  1200  ft.  between  the 
basal  quartzite  on  the  south  of  the  fault  and  of  that  resting  upon  the 
granite  on  the  north. 

The  vein,  which  fills  the  fault-fissure,  is  followed  on  the  south  side 
by  an  intrusion  of  fine-grained  green  diabase,  varying  in  width  up  to  70  ft. 


OVERHAND    STOPING    WITH   SHRINKAGE    AND    DELAYED    FILLING       159 

The  Coronado  oreshoot  is  approximately  2000  ft.  long  and  will 
average  35  ft.  in  width.  It  has  been  opened  by  a  three-compartment 
shaft  to  a  depth  of  700  ft.  The  vein  is  practically  vertical.  The  north 
or  foot-wall  is  of  slightly  altered  granite  and  the  south  or  hanging-wall 
is  of  quartzite  to  a  depth  of  150  ft.,  below  which  the  vein  enters  the 
granite  fissure.  The  zone  of  sulphide  ore  is  reached  at  a  depth  of  250  to 
300  ft.;  above  this  level,  small  bodies  of  oxidized  ore  have  been  found. 

The  ore  of  the  sulphide  zone  is  chalcocite,  in  some  places  entirely 
replacing  and  in  others  forming  a  coating  on  pyrite  and  chalcopyrite. 
The  gangue  consists  of  crushed  and  altered  granite  and  diabase;  in  this 
respect  the  vein  differs  from  most  of  the  others  of  the  district.  Horses 
of  granite  are  occasionally  found  in  the  vein,  the  outer  shells  of  which 
will  be  typical  ore,  gradually  merging  to  an  interior  of  slightly  altered 
granite,  showing  no  line  of  demarcation. 

MINING  METHODS 

The  orebody  is  contained  between  walls  of  granite,  the  foot-wall  is 
exceedingly  hard  and  the  hanging  wall  is  hard,  though  liable  to  slab  off 
in  large  pieces.  The  greater  par  tof  the  ore  is  of  medium  hardness  and, 
not  being  frozen  to  either  wall,  parts  readily  from  them.  A  back  of  ore 
will  generally  stand  well  without  support  if  properly  arched;  it  is  advisa- 
ble, however,  to  work  it  out  rapidly  to  prevent  "air  slaking/'  No 
sudden  change  in  the  width  of  the  orebody  has  been  found,  and  no  sul- 
phides occur  in  the  walls  as  is  often  the  case  in  the  porphyry  deposits. 

The  system  of  Example  22  was  formerly  employed  in  hard  ore,  when 
sufficiently  close  to  surface  to  allow  of  waste  filling  being  easily  obtained. 

The  expense  of  breaking  and  leveling  waste  for  each  slice  and  the 
exposing  of  unskilled  laborers  under  a  high  roof,  were  the  vital  objections 
to  the  continuance  of  this  system.  These  defects  were  overcome  in  the 
present  system  in  which  the  shovelers  work  outside  the  stope,  the  miner 
is  kept  so  close  to  the  back  of  ore  as  to  allow  constant  scrutiny  and  the 
handling  of  waste  is  reduced  to  a  minimum. 

In  preparing  a  level  for  stoping  by  the  new  system,  two  methods  have 
been  employed.  In  the  first  method  shown  in  Figs.  77  and  78,  all  of  the 
ore  between  the  walls  is  removed  for  a  length  of  75  ft.  and  to  a  height  of 
20  ft.  This  space  after  being  floored  with  2-in.  plank  is  filled  with  waste 
from  the  old  stopes  above  to  within  5  ft.  of  the  back.  New  roadways 
are  now  driven  in  the  foot-  and  hanging-walls,  paralleling  the  vein  at  a 
distance  of  15  ft. 

Chute  raises  are  carried  up  at  intervals  of  25  ft.  along  these  roadways 
and  connected  to  the  stope  by  crosscuts.  The  broken  ore  from  the 
stope  runs  through  the  cross-cut;  the  grizzly  allows  the  finer  material 
to  pass  into  the  chute  and  the  larger  pieces  are  broken  by  a  laborer 


160 


MINING    WITHOUT   TIMBER 


stationed  on  the  grizzly.     The  chutes  are  all  connected  on  the  level  of  the 
grizzlies  by  small  drifts,  from  which  a  ladderway  extends  to  the  level. 
In  the  second  method  shown  in  Figs.  77,  79,  and  80,  the  sill  floor  of  the 


FIG.  77. — Vertical  section. 


FIG.  78. — Cross-section  of 
•  stope  at  A-B. 


FIG.  79. — Cross  section  of 
stope  at  C-D 


Temporary  Pillar  to  be  removed 

pillar  at         FIG.  80.  by  slicing  section 

section   E-F.  G-H. 

STOPING  SYSTEM  AT  CORONADO  MINE. 


stope  is  started  15  ft.  above  the  tramming  level.  This  level  is  in  the 
center  of  the  vein  and  is  timbered  two  sets  high  for  the  whole  length 
of  the  stope,  leaving  a  shell  of  ore  between  the  top  of  the  timber  and  the 


OVERHAND    STOPING    WITH    SHRINKAGE    AND    DELAYED    FILLING       161 

floor  of  the  stope  above.  On  each  alternate  side  of  the  upper  sets,  in- 
clined funnel-shaped  raises  communicate  with  the  stope,  the  floor  of 
which,  viewed  from  inside  the  stope,  consists  of  two  rows  of  hoppers. 
The  broken  ore  passing  through  these  openings  falls  upon  the  6x8-in. 
lagging  with  which  the  floor  of  the  upper  sets  is  lagged.  By  opening  the 
center  lagging,  the  ore  is  raked  into  the  cars  placed  beneath. 

The  cost  of  the  preparatory  work  is  less  in  this  system  than  in  Fig. 
78,  the  loading  of  the  cars  being  direct  is  cheaper,  but  the  rapid  cleaning 
out  of  the  stope, which  is  an  important  matter,  is  subject  to  more  delays. 

PILLARS 

Two  classes  of  pillars  are  employed  to  support  the  roof  and  walls, 
small  temporary  stoping  pillars  and  larger  pillars  to  be  removed  later  by 
top  slicing.  A  section  along  the  vein  in  Fig.  77,  shows  a  pillar  of  ore  30 
ft.  long,  a  stope  of  75  ft.,  a  temporary  pillar  of  10  ft.  in  length,  another 
stope  of  75  ft.,  and  again  a  pillar  30  ft.  in  length.  The  30-ft.  pillar  is 
provided  with  a  chute  and  ladderway,  from  which  drifts  at  intervals  of 
15  ft.  give  entrance  and  ventilation  to  the  stopes.  The  smaller  pillar  is 
10  ft.  long  by  the  width  of  the  vein  and  contains  a  ladderway  with 
small  drifts,  as  in  the  larger  pillar. 

OVERHAND  STOPING 

The  ore  is  broken  by  overhand  stoping,  Waugh  drills  being  used. 
The  stope  is  kept  full  of  broken  ore,  sufficient  only  being  drawn  to  leave 
a  working  space  between  the  floor  of  broken  ore  and  the  back  of  the  stope. 
Work  is  confined  almost  entirely  to  the  ends  of  the  stope  adjacent  to  the 
pillars  with  the  purpose  of  leaving  a  sag  or  belly  of  ore  hanging  between. 
This  eventually  breaks  down  by  its  own  weight  and  is  block-holed 
from  on  top.  In  an  eight-hour  shift,  of  which  two  hours  are  consumed 
in  blasting  and  picking  down,  each  machine  will  drill  from  90  to  110  ft. 
of  holes. 

Should  the  back  of  ore  turn  soft  and  render  it  inadvisable  to  work 
beneath,  the  ore  can  be  broken  down  underhand  by  connecting  drifts 
from  the  raises  and  breaking  the  floor  as  described  in  connection  with  the 
King  Mine  of  Example  29. 

The  levels  are  200  ft.  apart,  and  when  a  stope  is  within  15  ft.  of  an 
upper  level,  the  breaking  of  ore  ceases.  Two  raises  are  then  made  from 
the  roof  of  the  stope  beneath  the  waste  filling  of  the  level  above.  The 
small  temporary  pillar  is  broken  by  first  undercutting  and  then  blasting 
from  inside  the  ladderway.  The  roof  of  the  stope  is  now  carefully  dressed 
and  the  stope  emptied  of  its  broken  ore  as  rapidly  as  possible. 
11 


162  MINING  WITHOUT  TIMBER 

FILLING  AND  EXTRACTING  PILLARS 

The  waste  filling  from  the  level  above  is  now  allowed  to  run  into  the 
empty  stope  and,  when  full,  a  working  floor  is  leveled  off.  To  extract 
the  shell  of  ore  left,  square-set  timbering  is  employed.  As  the  ore  is 
removed  by  retreating  to  the  chutes  in  the  pillars,  the  sets  are  caved  and 
the  waste  allowed  to  follow. 

To  extract  the  pillar  left  between  the  stopes,  work  is  commenced  be- 
neath the  upper  level.  Square  sets  are  employed  and  a  mat  laid,  upon 
which  the  waste  is  caved.  The  ore  is  then  removed  from  beneath  the 
mat  by  descending  slices  11  ft.  thick,  using  posts  to  support  the  over- 
hanging mat,  as  in  Example  43. 

When  the  broken  ore  is  drawn  from  the  stope  by  chutes  in  the  walls, 
there  must  necessarily  be  a  "  hog  back  "  of  broken  ore  left  along  the  center 
of  the  stope.  This  is  removed  by  spiling  a  timbered  roadway  through 
it  and  withdrawing  the  broken  ore.  As  soon  as  the  waste  appears, 
another  set  of  spiling  is  blasted  out,  retreating  in  this  manner  to  each 
end  of  the  stope. 

Where  the  sill  floor  is  above  the  roadway  as  in  Figs.  79  and  80,  the 
stope  which  next  ascends  from  below  is  carried  up  to  this  level  and  the 
shell  of  ore  removed  as  before  described.  The  system  has  been  satis- 
factory and  has  resulted  in  a  substantial  reduction  in  the  cost  of  mining. 

EXAMPLE  31. — Los  PILARES  MINE,  NACOZARI,  MEXICO 
(See  also  Example  24.) 

Irregular  Lenses  in  Porphyry;  Slicing  and  Delayed  Filling. — Where 
this  system  is  employed  the  sill  flooring  of  a  stope  is  practically  the  same 
as  method  B  in  Example  24.  The  ore  extracted  by  this  system  of  stop- 
ing  has  a  better  grade  than  that  mined  in  Example  24,  and  for  that 
reason  no  chance  is  taken  of  mixing  it  with  the  waste  filling.  The 
ground  is  also  much  firmer,  allowing  a  large  stoping  area  without  danger 
of  caving  and  thus  losing  the  stope.  At  intervals  of  from  12  to  20  ft.  cross- 
cuts from  the  main  drift  (which  is  generally  on  the  wall)  are  driven  into 
the  ore  for  a  distance  of  20  ft.  from  the  center  of  the  main  drift.  If  the 
orebody  is  not  wide,  these  cross-cuts  will  suffice  to  draw  the  broken  ore 
of  the  stope,  as  there  will  not  be  too  much  space  between  the  far  side  of 
the  stope  and  the  end  of  such  a  cross-cut  which,  because  of  its  purpose, 
is  termed  a  shovel-way.  If  the  orebody  is  wide,  as  in  Fig.  65,  an  auxil- 
iary drift  d  is  driven  through  the  ore  at  approximately  right  angles  to  the 
pillars  and  located  about  two-thirds  the  distance  between  the  country 
wall  and  the  limit  of  the  ore.  Cross-cuts  g  are  then  driven  at  intervals  of 
from  12  to  20  ft.  from  this  auxiliary  drift  to  both  the  left  and  the  right,  15 
ft.  long,  and  after  leaving  a  15-ft.  pillar  between  the  side  of  the  drift 
and  the  end  of  the  cross-cuts,  the  remaining  area  of  the  stope  is  "silled" 


OVERHAND    STOPING    WITH    SHRINKAGE    AND    DELAYED    FILLING       163 


on  the  level  floor.  The  pillar. thus  left  forms  the  base  of  the  floor  arch 
over  the  drift,  and  is  pierced  by  the  cross-cuts  which  serve  as  shovelways. 
Turn  sheets  are  then  placed  in  the  main  drift  in  front  of  each  cross-cut 
and  from  6  ft.  to  8  ft.  of  track  laid  in  each  cross-cut.  A  platform  h  raised 
2  ft.  above  the  rails  is  then  placed  in  the  cross-cut  at  the  far  end  of  the 
track  and  a  flat  iron  sheet  placed  on  top  of  it,  at  the  same  time  raising  the 
roof  above  the  platform  about  1  ft.  This  track  arrangement  allows  the 
car  to  be  turned  into  the  shovelway,  leaving  the  main  track  open.  The 
broken  ore  runs  down  on  to  the  iron  sheet  over  the  platform  from  which 
it  is  shoveled,  thus  giving  the  car  man  an  easier  task  in  filling  his  car. 

Fifteen  feet  above  the  level  floor,  the  stope  is  cut  out  back  over  each 
drift,  leaving  it  protected  by  the  pillars  and  floor  arch  i  as  seen  in  the 
section  shown  in  Fig.  65,  but  thus  acquiring  the  whole  area  between  the 


FIG.  81. — Town  of  Los  Pilares. 

pillars  overhead  for  stoping.  From  the  top  of  the  arch  up,  the  stope  is 
worked  by  two  and  three  slices  being  carried  forward  at  the  same  time. 
The  broken  ore  accumulates  in  the  stope  and  only  enough  is  drawn  from 
below  through  the  shovelways  to  allow  the  miners  to  drill  by  standing  on 
ore.  Manways  /  are  carried  up  through  the  center  of  the  pillars  limit- 
ing the  stope  on  each  side.  From  one  manway  20  ft.  above  the  level  floor 
an  intermediate  cross-cut  is  driven  to  the  stope  and  at  each  succeeding 
20  ft.  similar  cross-cuts  are  driven.  From  the  pillar  manway  on  the  other 
side,  the  first  intermediate  cross-cut  to  the  stope  is  driven  30  feet  above 
the  level  floor,  and  others  at  intervals  of  20  ft.  above  it.  These  man- 
ways  and  their  connecting  cross-cuts  give  an  inlet  and  outlet  to  the  stope 
for  every  10  ft.  By  this  method,  a  stope  may  be  carried  up  100  ft.,  or 
even  200  or  300  ft.,  before  being  drawn  and  filled.  Again  a  stope  may  be 
worked  from  two  or  three  levels  at  the  same  time,  each  level  being  driven 
up  till  only  a  thickness  of  12  to  15  ft.  of  solid  ground  separates  two  stopes. 


164  MINING    WITHOUT   TIMBER 

At  this  point  the  uppermost  stope  is  drawn  through  its  shovelways,  after 
which  the  floor  arches,  etc.,  protecting  the  haulage  drift  are  shot  down. 
The  solid  ground  separating  the  two  stopes,  one  over  the  other,  is  then 
drilled  with  a  large  number  of  holes,  say  50  to  60,  from  the  top  of  the 
broken  ore,  the  holes  heavily  loaded  and  shot  down,  thus  making  the  two 
stopes  one.  By  this  method  all  the  ore  left  for  floor  arches,  etc.,  is 
eventually  recovered  except  that  on  the  lowest  level  worked. 

These  great  stopes  may  or  may  not  be  filled  with  waste  rock  soon 
after  drawing  off  the  ore.  When  the  mine  was  first  opened  they  were 
sometimes  left  standing  empty  for  months  without  accident.  In  recent 
years,  however,  the  tendency  is  to  fill  as  soon  as  possible  after  drawing  the 
stopes.  As  illustrative  of  the  standing  qualities  of  the  rock,  no  better 
example  can  be  cited  than  that  of  the  old  No.  1  stope  worked  out  during 
the  first  years  of  mining.  This  was  located  near  the  Pilares  shaft 
(Fig.  81),  was  100x100  ft.  in  floor  plan  and  was  worked  up  from  the 
400-ft.  level  clear  to  the  oxidized  capping,  a  vertical  distance  of  280  ft. 
Although  the  capping  was  here  only  25  to  30  ft.  thick,  the  empty  stope 
stood  for  18  months  without  caving  before  it  was  filled. 

Filling  the  Stopes. — In  filling  stopes  mined  by  this  system,  fill  holes 
are  run  direct  to  the  surface  over  the  stope  where  they  are  widened  out 
to  a  size  of  12x12  ft.  On  the  surface  this  12xl2-ft.  hole  is  then  further 
enlarged  to  a  roughly  funnel  shape  by  churning  holes  from  12  to  30  ft. 
in  depth  above  the  edge  of  the  fill  hole.  These  churn  drill  holes  are 
sprung  with  dynamite,  and  shot  with  black  powder,  the  rock  breaking 
from  su  ch  shots  falling  directly  into  the  stope.  By  having  the  fill  hole 
of  this  large  size,  it  is  seldom  choked  with  large  rock. 


CHAPTER  XIII 

OVERHAND  STOPING  WITH  SHRINKAGE  AND   SIMULTANE- 
OUS PILLAR-CAVING 

EXAMPLE    32. — MIAMI   MINE,    GLOBE    DISTRICT,    ARIZONA 
(See  also  Examples  21  and  42.) 

Irregular  Lenses  in  Porphyry. — Rill  Chutes  and  Slicing  of  Pillars. — 
At  present  the  main  prospecting  shaft,  No.  1,  has  been  sunk  to  a  depth 
of  720  ft.  and,  as  ore  was  encountered  at  220  ft.,  its  thickness  at  the  shaft 
is  at  least  500  ft.  The  area  of  the  orebody  (see  Fig.  81)  is  about  10  acres 
and  it  is  covered  by  60  to  250  ft.  of  porphyry  capping. 

Below  the  220-ft.  level  sub-levels  have  been  driven  at  25-ft.  vertical 
intervals  down  to  the  370-ft.  level,  and  on  these  sub-levels  the  orebody 
has  been  almost  completely  blocked  out  into  50-ft.  squares  by  drifts  7  ft. 
high  and  from  4  1/2  to  5  ft.  wide. 

The  50-ft.  interval  of  ore  between  the  370-ft.  level  and  the  420-foot, 
or  main-haulage,  level  has  been  left  solid  to  protect  the  haulage  level 
while  mining  the  ore  above  it.  Before  detailing  the  method  of  mining, 
a  word  as  to  the  character  of  the  ore. 

Superintendent  N.  O.  Lawton,  the  deviser  of  this  mining  system, 
says:  "The  enriched  schist  of  the  spheroidal-shaped  orebody  has  been 
greatly  fractured,  crushed,  and  later  softened  or  altered  by  percolating 
water,  so  that  in  its  present  form  it  is  quite  easily  drilled  and  broken. 
The  fractured  condition  will  facilitate  mining  by  causing  the  ore  to 
break  into  pieces  easily  handled  by  the  shovel  or  run  in  chutes.  Little 
of  the  ore  will  break  in  so  large  pieces  as  to  necessitate  rebreaking  or 
block-holing  to  avoid  choking  the  chutes.  Most  of  the  ore  caves  easily, 
so  that  to  avoid  timbering  except  at  soft  places,  the  levels  are  driven  the 
narrow  width  of  4  1/2  to  5  ft.  by  7  ft.  high,  with  the  roof  carefully  arched." 

System  of  Mining. — It  is  apparent  from  the  above  description  of  the 
ore  that,  because  of  its  soft  crushed  nature,  which  causes  it  to  quickly 
cave  wherever  excavations  of  any  width  are  left  untimbered,  a  system 
of  caving,  practical  where  the  ore  is  hard  and  stands  well,  would  here  be 
unsuitable. 

The  accompanying  illustrations,  Figs.  82  to  85,  will  make  clear  the  fol- 
lowing description.  Fig.  82  is  a  plan  of  the  orebody  showing  method  of 
rectangular  system  for  tramming  levels  located  50  ft.  below  floor  of 
stopes.  Fig.  83  is  a  cross-section  through  rooms  and  pillars.  Fig.  84 

165 


160 


MINING    WITHOUT    TIMBER 


is  a  plan  of  first  mining  level,  showing  method  of  cutting-out  room 
preparatory  to  stoping.  Fig.  85  is  a  detail  plan  of  sub-levels,  showing 
method  of  stoping  rooms. 

On  the  first  main-haulage  level,  the  420-ft.  haulage  drifts  have  been 
driven  spaced  on  50-ft.  centers,  as  shown  in  Fig.  82.  The  ground  is 
to  be  excavated  by  a  series  of  rooms  60  ft.  wide  alternating  with 
40-ft.  pillars  as  best  understood  by  reference  to  Fig.  83.  Every  alter- 
nate haulage  drift  is  to  be  provided  with  a  mill  hole  to  draw  the 
broken  ore*  from  the  rooms  while  the  other  drifts  will  come  beneath  the 


FIG.  82.  First  haulage  level,  Miami  mine. 

pillars.  From  these  haulage  drifts  4x5  ft.  raises  e  have  been  put  up  so  as 
to  come  in  the  center  of  both  the  rooms  and  pillars.  This  spacing  ar- 
rangement has  necessitated  the  carrying  of  the  pillar  25  ft.  thick  on  one 
side  of  a  pillar  drift,  and  only  15  ft.  on  the  other,  as  shown  in  the  plan 
Fig.  84.  The  haulage  drifts  beneath  the  rooms  have  chutes  spaced  on 
25-ft.  centers,  while  the  drifts  beneath  the  pillars  have  chutes  every 
50  ft.  It  will  be  noted  on  Fig.  83  that  the  raises  to  the  rooms  are 
branched  from  a  point  25  ft.  above  the  haulage  level  in  order  to  make 
them  effective  in  drawing  the  broken  ore  all  the  way  across  the  room. 
As  previously  stated,  the  orebody  has  been  already  completely  blocked 
out  preparatory  to  starting  caving  in  the  near  future. 


OVEKHAND    STOPING    WITH    SHRINKAGE    AND    PILLAK-CAVING 


167 


Operations  will  start  on  the  first  mining  level,  Fig.  83,  50  ft.  above  the 
haulage  level.  The  top  line  of  Figs.  84  or  85,  represents  the  limit  of  com- 
mercial ore.  Mining  will  commence  at  this  limiting  line  and  retreat  from 
it.  Starting  then  at  the  end  of  the  drift  cross-cuts  will  be  driven  both  ways 
along  the  limiting  line  to  the  pillar  lines.  These  cross-cuts  will  be  carried 
as  wide  and  high  as  found  practical,  say  7  ft.  high  by  8  ft.  wide.  The 
roof  of  these  cross-cuts  will  then  be  drilled  with  upper  holes,  using  ham- 
mer-drill stopers,  and  shot  down  to  a  height  of  7  ft.  more,  making  the  total 
height  of  the  cross-cuts  14  ft.  Then  starting  again  at  the  drift  a  slice  or 
drivage  will  be  -taken  each  way  in  a  similar  manner,  breaking  to  the  pre- 
vious excavation  and  then  shooting  to  obtain  the  height  of  14  ft.  In 


25'  50'  75'  100' 

FIG.  83. — Cross-section  of  stope,  Miami  mine. 

this  way,  retreating  slice  by  slice  from  the  limiting  line  the  ore  above  the 
floor  of  the  room  is  broken  to  a  height  of  14  ft.,  although  at  no  point  do 
the  miners  work  under  a  roof  higher  than  7  ft.  The  ore  thus  broken 
will  fill  the  branched  raises  of  Fig.  83,  previously  driven;  but  only  as 
much  ore  as  necessary  will  be  drawn,  as  it  is  desirable  that  it  shall  pack 
under  the  solid  back  as  closely  as  possible,  thus  supporting  its  weight. 

The  miners  are  then  transferred  to  the  first  submining  level  located 
25  ft.  above.  As  14  ft.  of  ore  has  been  broken  below,  the  thickness  of 
solid  ground  underfoot  is  now  but  11  'ft.  Reference  to  Fig.  83  will 
make  clear  the  method  of  procedure  to  break  up  this  11  ft.  Starting, 
at  a  central  raise  down  holes  are  placed  all  around  it  and  shot  through 
to  the  broken  ore  in  the  room  beneath.  Simultaneously  work  is  started 
at  the  next  raise  and  the  floor  likewise  broken  down.  Drill  men  then 


168 


MINING    WITHOUT   TIMBER 


drill  the  floor  and  roof  of  the  drift  which  connects  the  two  raises  blasting 
them  both  simultaneously  in  4-  to  5-foot  sections,  retreating,  from  the 
holes  started  at  the  raises,  toward  each  other  till  they  meet  about  in  the 


50  100 

FIG.  84. — Plan  of  first  mining  level,  showing  chutes,  Miami  mine. 

center  of  the  drift.  This  work  breaks  the  floor  of  the  drift  through  into 
the  room  full  of  broken  ore  below,  and  breaks  up  its  roof  to  a  height  of 
14  ft.  The  drill  men  now  start  again  at  the  raises,  and  leaving  in  the 
solid  floor  to  work  on,  slice  back  the  walls  for  a  width  of  about  7  ft.  but 


FIG.  85. — Plan  showing  rooms  during  sloping,  Miami  mine. 

only  7  ft.  high,  thus  paralleling  the  slot  broken  through  to  the  room 
below  and  again  meeting  in  the  middle  of  the  block  as  before.  The  floor 
and  roof  are  now  simultaneously  drilled  and  shot,  breaking  through  into 


OVERHAND    STOPING    WITH   SHRINKAGE    AND    PILLAR-CAVING          169 

the  room  below,  and  raising  the  roof  to  a  height  of  14  ft.  In  this  manner 
slice  after  slice  is  worked  back  parallel  with  the  original  drift  until  the 
pillar-line  limits  are  reached,  thus  breaking  up  the  entire  25-ft.  block 
below  the  first  submining  level,  and  likewise  demolishing  the  lower  14  ft. 
of  the  block  below  the  second  submining  level.  The  miners  are  then 
transferred  to  the  second  submining  level,  and  the  breaking  process 
repeated  from  sublevel  to  sublevel  until  the  leached  capping  is 
reached. 

In  spite  of  the  fact  that  the  soft,  crushed  nature  of  the  rock  makes 
it  a  bad  roof  to  work  under,  it  will  be  noted  that  the  method  always 
provides  safety  by  keeping  the  height  of  the  roof  under  which  miners 
work  at  only  7  ft.  and  affording  ample  ways  of  ingress  and  egress.  Owing 
to  the  shattered  condition  of  the  rock  there  should  be  no  difficulty  in 
breaking  it  fine  enough  to  avoid  choking  of  the  chutes. 

For  placing  the  side  and  down  holes,  light  2  1/4-in.  one-man  piston 
drills  will  probably  be  used,  and  for  uppers  the  air-hammer  drill. 

It  will  be  readily  apparent  that  quite  a  number  of  men  can  be  worked 
in  a  room  when  it  is  once  started  from  several  points,  and  to  that  end 
the  compressed-air  pipe  system  now  being  installed  has  been  designed  to 
furnish  air  to  a  large  number  of  drills  per  level  without  reducing  the 
desired  pressure.  The  main  air  pipe  is  10  in.  in  diameter  down  the  shaft, 
and  from  it  two  8-in.  diameter  pipes  are  branched  off  at  the  main  haulage 
levels,  the  420-  and  570-ft.  On  each  haulage  level  the  8-in.  pipe  is  hooped 
completely  about  the  rectangle  of  Fig.  82.  Each  of  the  haulage  drifts  I 
branching  from  the  rectangle  is  to  have  a  4-in.  main  into  which  2  1/2-in. 
pipes  are  tapped  and  carried  up  the  pillar  raises  to  the  various  sublevels 
where  1  1/2-in.  pipes  connect  with  the  drills.  By  this  system  of  piping, 
an  ample  supply  of  air  at  the  proper  pressure  is  provided. 

Drawing  the  Broken  Ore. — As  the  ore  runs  about  12  cubic  feet  to  the 
ton  solid  and  about  16  to  17  cubic  ft.  broken,  it  will  be  understood  that 
as  the  breaking  up  of  the  rooms  proceeds,  enough  of  the  ore  is  drawn 
through  the  chutes  to  allow  a  sufficient  space  before  blasting  for  a  free 
breaking.  Yet  at  all  times  the  top  of  the  broken  material  is  kept  close 
under  the  solid  back  in  order  to  prevent  the  falling  of  the  large  masses 
into  it.  Such  would  have  to  be  block  holed  under  a  roof  of  dangerous 
height  to  prevent  choking  of  the  chutes.  After  two  adjoining  rooms 
have  been  broken  the  length  of  the  ore  body,  or,  say,  200  to  300  ft.  back 
from  the  limiting  line  of  Fig.  82,  the  intermediate  pillar  may  be  mined 
by  the  method  of  slicing  of  Example  43,  as  shown  at  the  top  of  Fig.  76 
(6).  It  will,  of  course,  be  understood  that  the  irregular  dome-shaped 
masses  of  comparatively  small  horizontal  area  are  best  mined  by  this 
same  slicing  method,  before  the  rooms  are  broken  up  to  the  level  where 
the  width  of  the  ore  is  fairly  uniform  across  the  whole  ore  body.  This 
slicing  system  breaks  the  capping  to  the  surface  and  gets  its  weight  on 


170  MINING    WITHOUT    TIMBEK 

the  ore  body  before  the  rooms  reach  the  top  submining  level  shown  in 
Fig.  82. 

With  the  rooms  on  each  side  of  the  pillar  broken  up  to  the  top  sub- 
mining level,  the  first  slice  of  the  pillar  is  mined  until  the  broken  ore  is. 
encountered.  The  level  of  the  broken  ore  in  the  rooms  is  then  lowered  to 
that  of  the  top  of  the  solid  pillar,  thus  allowing  the  mat  of  timbers  and 
broken  materials  n  to  come  down  and  bringing  the  weight  of  the  roof 
on  the  timbers  above  the  pillar,  thus  crushing  them  flat  to  form  the  mat. 
A  second  slice  of  the  pillar  is  then  mined,  the  ore  in  the  rooms  drawn 
down  to  the  new  top  of  the  pillar,  and  so  on.  Eventually  the  ore  is  all 
broken  and  drawn  down  to  the  first  mining  level.  The  main  haulage 
system  of  the  mine,  which  up  to  this  time  has  been  on  the  420-ft.  level, 
will  then  be  transferred  to  the  570-ft.  level,  and  mining  will  proceed 
above  it,  starting  on  the  first  mining  level  50  ft.  above,  or  the  520-ft. 
level.  In  this  way  the  50-ft.  block  between  the  420-ft.  and  the  370-ft. 
level  will  ultimately  be  mined. 

By  getting  a  thick  mat  of  timbers  on  the  top  submining  level  before 
the  drawing  of  the  rooms  commences,  and  as  the  capping  comes  down 
keeping  it  at  as  uniform  a  level  as  possible  by  drawing  the  chutes  uni- 
formly, it  is  expected  that  little  ore  will  be  lost  by  mixing  with  waste. 

Tramming. — The  drifts  on  the  420-ft.  haulage  level  have  been  laid 
with  30-lb  rails  set  to  a  24-in.  gauge.  The  ore  will  be  drawn  from  the 
chutes  into  cars  of  2  1/2  tons  capacity  and  hauled  in  trains  to  the  shaft 
by  electric  locomotives.  At  the  shaft  the  cars  will  discharge  into  a 
700-ton  pocket,  below  which  will  be  a  skip  filler  holding  exactly  a  skip 
load.  The  skips  will  be  of  71/2  tons  capacity.  With  a  skip  placed 
beneath  the  filler,  a  lever  will  discharge  the  filler  into  the  skip. 

Summary  of  Mining  Conditions. — Thanks  to  the  compact  nature  of 
the  Miami  ore  body  and  the  character  of  the  ground,  ore  can  be  mined 
with  the  use  of  little  or  no  timber  and  a  minimum  of  explosive.  A  force 
of  but  comparatively  few  men  can  break  and  handle  enough  ore  to  main- 
tain a  steady  output  of  2000  tons  per  day.  With  the  ground  thoroughly 
blocked  out  and  with  the  present  equipment  as  it  is,  the  doubling  of  the 
output  would  be  simply  a  matter  of  increasing  the  number  of  miners. 

The  management  has  established  the  cost  of  mining  the  present  ore 
body  at  $1.25  per  ton,  but  the  average  cost  should  eventually  closely 
approximate  $1  per  ton,  including  hoisting  to  the  surface.  The  method 
of  mining  seems  well  adapted  to  insure  both  the  safety  of  the  miners  and 
the  extraction  of  at  least  85  per  cent,  of  the  ore  body.  Exclusive  of 
mining  losses,  therefore,  the  present  development  justifies  an  expec- 
tation of  hoisting  at  least  14,000,000  tons,  which  is  said  to  average  2.75 
per  cent,  copper  as  determined  by  systematic  churn-drill  prospecting. 

As  the  development  plan  for  this  system  is  similar  to  that  of  several 
others,  like  the  room-caving  of  Example  46,  or  the  block-caving  of 


OVERHAND    STOPING    WITH    SHRINKAGE    AND    PILLAR-CAVING  171 

Example  42,  the  st oping  could  be  altered  to  follow  one  of  these  similar 
systems  at  any  time  that  it  might  seem  desirable. 

EXAMPLE  33. — BOSTON  CONSOLIDATED  MINE,   BINGHAM,  UTAH 
(See  also  Examples  3,  37,  41  and  43.) 

Irregular  Lenses  in  Porphyry.  Block  Caving  of  Pillars. — A  caving 
system  has  been  devised  and  adopted  here.  Briefly,  this  consists  in 
weakening  the  block  of  ore  by  means  of  a  series  of  ore-filled  rooms,  and 
then,  when  the  remaining  pillars  are  shattered,  the  ore  is  drawn  evenly 
from  under  a  large  area  of  capping  and  the  surface  allowed  to  settle 
gradually. 

The  mine  is  opened  up  by  two  levels,  one  about  200  ft.  below  the 
steam-shovel  workings  near  the  top  of  the  hill  and  the  other,  the  main 
haulage  level,  150  ft.  below  the  first.  On  this  lower  level  there  are  two 
main  haulage  drifts.  From  these  a  system  of  parallel  side  drifts  are  turnd 
off  at  intervals  of  120  ft.  At  distances  of  200  ft.  along  these  drifts,  raises 
5  ft.  square,  inclined  at  an  angle  of  60  deg.,  are  driven  to  connect  with 
drifts  on  the  level  above,  and  from  these  main  raises,  or  chutes,  as  they 
become  later,  a  series  of  branch  raises  or  chutes  are  driven  so  that  the 
collecting  of  ore  may  be  concentrated  as  much  as  possible.  These  raises 
are  equipped  with  air-operated  sector  gates,  which  have  24-in.  openings, 
and  work  admirably  as  a  train  of  ten  8-ton  cars  can  be  loaded  in  4  min. 

The  tops  of  these  raises  and  branch  raises  hole  into  the  center  of  the 
drifts  on  the  level  above,  so  that  they  have  to  be  fenced  off  by  means  of 
guard  rails  nailed  to  four  upright  sprags.  A  series  of  knotted  ropes, 
hanging  from  the  guard  rails  like  the  low-bridge  signals  on  railroad 
tracks,  permit  the  placing  of  the  guard  rails  high  enough  not  to  interfere 
.with  the  dumping  of  a  car.  This  arrangement  prevents  any  one  from 
walking  into  the  chutes. 

BLOCKING  OUT  THE  ROOMS 

On  the  upper  level  (as  in  Fig.  86)  the  orebody  is  blocked  out  by  a 
series  of  drifts  60  ft.  apart  and  a  series  of  cross  drifts  400  ft.  apart,  since 
that  is  the  length  of  the  room  used  wherever  the  size  of  the  orebody 
permits.  These  drifts  are  6  1/2x7  1/2  ft.  in  the  clear  and  are  driven  on 
contract  at  $2  to  $2.25  per  foot,  making  their  total  cost  $3  to  $3.25  per 
foot.  In  driving  these  drifts  a  5-ft.  round  is  drilled  each  shift  by  one 
man  using  a  2  1/2-in.  Sullivan  piston  drill,  but  rarely  does  the  round 
break  over  4  ft.  in  the  clear.  The  powder  consumption  is  about  2  Ib.  to 
the  foot.  These  drifts  in  the  future  pillars  are  laid  with  double  track,  so 
as  to  permit  rapid  tramming. 

At  intervals  of  30  ft.  along  the   pillar  drifts,  drifts  30  ft.  long   are 


172 


MINING    WITHOUT   TIMBER 


driven  in  both  directions.  The  mouths  of  the  drifts  are  staggered,  so  as 
not  to  come  opposite  to  each  other,  thus  possibly  weakening  the  pillar  too 
much.  In  addition  the  tracks  in  these  short  drifts  are  arranged  so  that 
the  cars  are  run  out  to  the  nearer  cross  drift.  The  ends  of  these  short 
drifts  are  then  connected  with  each  other  by  means  of  drifts  parallel  to 
the  pillar  drifts.  The  short  drifts  at  the  end  of  each  pillar  are  not  driven 
within  15  ft.  of  the  main  cross  drifts,  in  order  not  to  weaken  their 
pillars. 


y 


---60--- 

FIG.  86  — Stoping  system,  Boston  Con.  mine. 


As  soon  as  these  short  drifts  are  all  connected,  this  central  drift,  which 
marks  the  middle  of  the  future  room,  is  slabbed  out  by  means  of  two  flat 
holes,  one  near  the  floor  and  the  other  near  the  roof.  These  holes  are 
drilled  from  16  to  18  ft.  deep  and  are  parallel  to  the  drift.  By  blasting 
two  rows  of  holes  on  each  side  of  the  room  drift,  the  room  is  enlarged  to 
a  width  of  30  ft.  Model  9  Water  Leyner  drills  are  used,  and  in  drilling 
these  long  holes  a  2  1/2-in.  starter  and  a  1  1/2-in.  finisher  bit  is  employed. 
A  1-in.  water  pipe  is  carried  along  the  pillar  drift  to  water  these  drills. 

In  alternate  pillars,  at  the  same-  time  that  the  room  drifts  are  being 
driven,  raises  are  put  up  at  the  middle  of  the  pillar  drift  and  also  near 
each  end  where  the  pillar  drift  meets  the  main  cross  drifts.  These  serve 


OVERHAND    STOPING    WITH    SHRINKAGE    AND    PILLAR-CAVING          173 

as  future  manways  in  mining  the  rooms.  They  are  driven  merely  large 
enough  for  a  ladder  and  an  air  pipe. 

As  soon  as  the  room  has  been  enlarged  to  a  width  of  30  ft.  and  the  ore 
mucked  out,  a  chute  having  a  mouth  24  in.  wide  is  built  across  the  end  of 
each  short  cross  drift.  These  chutes  are  used  in  drawing  off  the  excess 
of  broken  ore  from  the  room.  Miners  then  begin  to  drill  the  roof  of  the 
room  full  of  uppers,  using  Leyner  air-hammer  drills.  These  holes  are 
drilled  so  as  to  look  forward  about  75  deg.  and  toeing  out  to  the  sides  so 
as  to  keep  the  walls  of  the  room  vertical.  These  holes  are  drilled  about 
7  ft.  deep  and  in  rows  of  5  holes  across  the  stope,  these  rows  being  8  to  9  ft. 
apart.  It  takes  about  5  drills  to  a  hole,  the  starters  having  a  1  1/2-in. 
bit  and  the  finishers  a  1-in.  bit.  These  Leyner  hammer  drills  use  solid 
steel  and  drill  from  100  to  120  ft.  in  a  shaft,  the  air  pressure  being  only 
80  Ib.  per  sq.  in.  at  the  receiver.  This  low  pressure  is  used  because  with 
a  higher  pressure  the  bits  cut  too  fast  to  permit  them  to  turn  freely. 

These  holes  are  not  loaded  until  half  the  length  of  the  stope  has  been 
drilled.  Then  they  are  loaded  with  about  4  sticks  of  30  per  cent.  Her- 
cules dynamite  and  blasted  all  at  once.  About  0.45  Ib.  of  dynamite  is 
required  per  ton  of  ore  broken.  One  man,  by  using  a  nicked  fuse,  spits 
about  50  holes,  a  7-ft.  fuse  that  burns  at  about  45  sec.  per  foot  being  used. 
Thus  the  stope  is  carried  up  in  alternate  halves. 

The  excess  ore,  which  amounts  to  about  40  per  cent,  of  the  total  ore 
broken,  is  drawn  off  before  each  blast,  so  that  there  is  12  ft.  of  open  space 
below  the  roof  when  the  holes  are  spit.  The  ore  at  first  was  drawn  off 
at  only  one  side,  but  this  left  the  pile  slanting  toward  that  side  and  neces- 
sitated leveling  it  off  each  time  before  drilling  began.  To  avoid  this  the 
tapping  drifts  are  driven  frpm  each  side  and  the  ore  drawn  off  on  both 
sides  so  that  it  keeps  a  fairly  level  surface  for  the  men  to  work  upon. 

Three  manways,  one  at  each  end  and  one  in  the  center  of  the  side  of 
the  pillar  having  the  raise  in  it,  are  carried  up  through  the  ore.  These 
are  "made  by  blasting  out  a  triangular  notch  in  the  pillar  and  cribbing  it 
off  on  three  sides.  This  manway  is  about  3x3  ft.  in  size  and  is  merely 
large  enough  to  allow  a  man  to  pass  through  it  after  the  air  pipe  has  been 
put  in.  The  men  climb  the  cribbing  of  split  timber,  a  vertical  pole  being 
nailed  to  the  cribbing  for  the  men  to  hold  to  when  climbing.  These 
manways  are  carried  up  through  the  ore  for  only  50  ft.,  since  at  every 
50  ft.  a  drift  is  driven  to  the  stope  from  the  three  raises  in  the  pillar. 

The  ore  is  drawn  off  in  cars  having  a  capacity  of  28  cu.  ft.  and  is  run 
by  two  men  to  the  main  chutes.  Only  4  or  5  men  work  in  a  stope,  and 
one  of  these  is  busy  all  the  time  picking  down  the  back  of  the  room  and 
the  roofs  of  the  different  drifts  in  the  pillar  connecting  with  that  room. 
The  rooms  are  carried  up  until  the  ore  becomes  too  poor  to  pay — at  pres- 
ent when  it  carries  about  1.45  per  cent,  copper.  The  room  is  then  aban- 
doned, and  as  soon  as  the  stope  on  the  other  side  of  the  pillar  is  completed, 


174  MINING    WITHOUT    TIMBER 

the  air  pipes  and  tracks  in  the  raises  and  drifts  in  the  pillar  between  the 
two  stopes  are  removed. 

CAVING  THE  ORE 

As  the  area  undermined  by  rooms  increases,  the  roof  gradually  settles; 
but  as  the  top  of  the  broken  ore  in  the  rooms  is  within  six  feet  of  the  roof 
when  the  rooms  are  abandoned,  the  capping  cannot  drop  far.  This 
undercutting  throws  weight  on  the  pillars,  and  after  an  area  200  by  400 
ft.  has  been  undermined,  they  begin  to  crush  without  any  further  weak- 
ening. At  present  five  stopes  are  being  worked  on  the  upper,  and  three 
on  the  lower  level. 

After  the  orebody  on  the  upper  level  has  been  undercut  by  rooms, 
mining  on  the  lower  level  begins  under  that  area.  On  the  lower  level 
the  rooms  are  placed  so  as  to  come  directly  under  the  pillars  on*  the  level 
above.  These  rooms  are  mined  in  the  same  manner  as  on  the  level  above, 
with  the  exception  that  their  floors  are  30  ft.  above  the  level,  so  as  to  leave 
a  pillar  30  ft.  thick  to  protect  the  main  transportation  drifts.  This 
necessitates  the  driving,  from  the  main  level,  of  inclined  raises,  30  ft. 
apart,  to  tap  the  rooms  on  each  side  so  as  to  draw  off  the  excess  ore.  The 
manways  in  the  pillars  are  also  dispensed  with  in  the  lower  blocks  of 
ground,  for  it  has  been  found,  owing  to  the  weight  on  the  lower  pillars, 
to  be  cheaper  to  drive  drifts  to  the  different  rooms  at  intervals  of  50  ft. 
vertically  from  a  raise  placed  in  a  part  of  the  lower  level  not  being  under- 
mined than  to  maintain  the  manways  in  the  pillars. 

After  the  whole  of  the  orebody  has  been  cut  up  by  rooms  and  the 
lower  pillars  have  been  weakened  by  raises  put  up  to  tap  them,  drawing- 
will  begin  throughout  the  whole  orebody,  the  ore  being  removed  evenly 
under  the  area  so  that  the  capping  will  settle  regularly.  Thus  the  mixing 
of  ore  and  capping  will  be  prevented  as  much  as  possible.  When  this  is 
completed,  the  ore  below  this  level  and  the  pillar  above  it  will  be  mined 
by  a  similar  method  of  undercutting  by  means  of  rooms. 

The  company  is  mining  2600  dry  tons  of  ore  a  day  at  its  porphyry 
mine,  and  employs  351  men  underground  and  on  surface.  This  gives  an 
average  of  almost  7  tons  to  a  man  and  from  8  to  8  1/2  tons  to  a  man 
employed  underground.  This  is  remarkable  when  one  considers  that 
only  ore  broken  in  the  rooms  is  being  taken  out  at  present. 

The  wages  of  the  men  per  8-hr,  shift  are:  Machinerren  and  timbermen, 
$3;  helpers  and  muckers,  $2.50;  trainmen,  $3;  helpers,  $2.50;  blacksmiths, 
$4;  toolsharpeners,  $3.50;  shift  boss,  $4. 

The  cost  of  driving  the  large  9x9-ft.  haulage  drifts  is  about  $8  per 
foot,  as  the  cost  of  such  a  drift  1600  ft.  long,  one-third  timbered,  was 
$8.08  per  foot,  inclusive  of  the  cost  of  laying  track  and  of  putting  in  the 
trolley  wire.  The  cost  of  driving  a  6  1  /2x7-ft.  drift  is  for  labor  (contract), 
$2  to  $2.25  per  foot,  making  the  total  $3  to  $3.25.  For  driving  5x5-ft, 


OVERHAND    STOPING    WITH    SHRINKAGE    AND    PILLAR-CAVING  175 

raises  the  cost  is,  for  labor  (contract),  $1.75,  making  the  cost  per  foot 
for  raises  180  to  200  ft.  high  $2.25  to  $2.50  per  foot. 

At  present  the  cost  of  mining  the  ore  is  44  cents  per  ton. 
The  cost  of  development  is  2  cents  per  ton  of  ore  developed  in  a 
room  and  a  pillar,  for  the  company  charges  this  at  the  rate  of  10 
cents  per  ton  to  the  ore,  that  is  drawn  off  from  a  room,  or  to  only  40 
per  cent,  of  the  total  ore  broken  in  a  room.  The  rooms  and  pillars  are 
equal  in  size.  Therefore,  disregarding  the  ore  removed  from  the  raises 
and  drifts  in  the  pillars,  which  work  comes  under  the  head  of  develop- 
ment, there  remains,  after  the  whole  area  is  caved,  80  per  cent,  of  the  ore 
that  needs  only  to  be  drawn  through  chutes  and  trammed  to  surface. 

At  present  it  costs  17  cents  a  ton  to  draw  the  ore  from  the  rooms,  so 
that  the  cost  of  drawing  the  60  per  cent,  left  in  the  rooms  and  pillars 
ought  not  to  cost  more  than  20  cents  per  ton.  Consequently  the  cost  of 
mining  figures  out  as  follows :  Forty  tons  of  ore  mined  at  44  cents  per  ton, 
$17.60;  160  tons  of  ore  mined  at  20  cents  per  ton,  $32;  development  cost 
for  40  tons  at  10  cents  per  ton,  $4.  Total  ^cost  of  extracting  200  tons, 
$53.60;  or  26.8  cents  per  ton.  The  cost  of  superintendence,  taxes,  etc., 
when  2600  tons  of  ore  are  mined  a  day,  will  amount  to  about  2  cents  per 
ton.  Allowing  a  factor  of  safety  of  25  per  cent.,  or  6  cents  per  ton,  to 
cover  unforeseen  difficulties  in  mining  the  lower  pillars  of  ore,  it  appears 
that  by  this  system  the  ultimate  cost  of  mining  a  block  of  ore  will  prob- 
ably approximate  35  cents  per  ton,  the  same  as  steam-shovel  mining. 

This  method  of  mining  is  quite  bold,  but  from  the  results  obtained  on 
the  small  area  already  caved,  it  appears  that  the  ore  breaks  in  a  fairly 
perpendicular  plane  to  surface.  This  will  greatly  decrease  the  tendency 
of  the  ore  and  capping  to  mix  when  it  becomes  necessary  to  cave  one 
section  of  the  orebody  before  an  adjacent  block  is  touched.  But  the 
main  difficulty  from  mixing  of  the  ore  and  capping  will  come  from 
unequal  settling  of  different  portions  of  the  block  that  is  being  caved. 
The  Bingham  porphyry  ore,  from  the  nature  of  its  formation,  is  much 
broken  up  by  small  fracture  seams.  Owing  to  this  fractured  nature  of 
the  orebody  on  caving,  it  breaks  up  into  fairly  fine  ore,  and  so  there  will 
be  few  large  boulders  to  block  the  chutes  or  to  cause  the  ore  to  hang  up 
above  the  chute  mouth  and  form  a  grizzly  through  which  only  fine  ore 
could  pass,  as  would  be  the  case  if  large  masses  followed  down  that  were 
crushed  but  little  by  the  weight  thrown  upon  them  by  the  undermining 
of  the  block.  It  therefore  appears  that,  if  a  system  of  recording  the 
approximate  tonnage  drawn  from  each  chute  is  used,  there  should  be  no 
great  difficulty  in  drawing  the  ore  evenly  from  under  the  capping  in  each 
block,  and  consequently  no  trouble  in  causing  the  capping  to  follow 
evenly  af  er  the  ore  with  little  intermixing  of  capping. 

The  advantages  of  this  method  of  caving  are  many:  The  ore  is  broken 
in  large  rooms;  the  method  of  stoping  is  adapted  to  the  use  of  air-hammer 


176  MINING    WITHOUT   TIMBER 

drills;  the  blasting  is  done  over  large  areas  so  that  as  little  time  is  con- 
sumed by  blasting  as  is  possible,  and  the  air  vitiated  only  between  shifts; 
the  work  is  done  systematically  throughout  an  area,  so  that  one  can  tell 
when  all  portions  of  the  block  have  been  reached;  the  ore  is  only  dropped 
a  few  feet  at  a  time  and  over  large  areas,  so  as  to  diminish  the  amount  of 
ore  mixed  with  capping;  the  ore  is  drawn  systematically  so  that  there  is 
little  danger  of  leaving  ore  behind;  the  amount  of  drifting  and  raising 
required  in  mining  a  block  of  ore  is  very  small;  the  gathering  of  the 
ore  is  rendered  cheap  by  concentrating  it  at  a  few  chutes  equipped 
with  easily  and  rapidly  operated  gates. 

Ground  to  be  adapted  to  this  method  must  be  weak  enough  to  cave 
readily,  and  yet  strong  enough  to  stand  in  rooms  20  or  30  ft.  wide.  The 
ore  must  break  up  into  small  chunks  when  the  weight  of  the  capping 
causes  the  pillars  to  cave,  or  otherwise  difficulty  will  be  experienced  in 
drawing  ore  from  the  caved  area.  Apparently  there  should  not  be  much 
difficulty  in  drawing  the  ore  on  the  lower  level,  as  the  pillar  30  ft.  thick 
ought  to  be  amply  strong  enough  to  protect  the  haulage  drifts. 

EXAMPLE  34. — DULUTH  MINE,  CANANEA,  MEXICO 
(See  also  Examples  6,  18  and  45.) 

Irregular  Lenses  in  Porphyry;  Block-caving  of  Pillars. — Pillar-caving 
is  a  combination  of  overhand  stoping  on  ore  and  a  caving  system.  As 
is  necessary  is  nearly  all  caving  methods,  the  first  step  is  to  prospect  and 
thoroughly  outline  the  orebody  by  means  of  drifts  and  raises.  Fig.  87 
shows  an  orebody  on  the  200  level  which  extends  above  the  100  level. 

After  sufficient  prospecting  work  has  been  done,  the  size  of  the  sec- 
tions to  be  mined  and  the  pillars  of  oje  to  be  left  were  decided  upon. 
Pillars  are  usually  about  50  ft.  wide,  with  sections  from  75  to  100  ft. 
wide  extending  across  the  body.  Because  of  the  irregularity  of  the 
upper  portions  of  these  bodies  it  is  necessary  that  they  be  mined  by 
means  of  square  sets  in  order  to  follow  rich  stringers. 

At  the  Cananea-Duluth  the  orebody  is  mined  by  square  sets  from 
the  100  level  to  the  top  of  the  ore.  These  sets  are  then  all  removed  and 
the  pillar-caving  system  proper  begins.  In  the  meantime  the  section  to 
be  mined  is  .blocked  out  on  the  200  level  by  means  of  drifts  and  regular 
square-set  raises  are  put  in  at  intervals,  as  shown  in  Fig.  87.  The  sill 
raise  set  is  8  ft.  5  in.  high  and  the  second  set  is  7  ft.  4  in.,  making  prac- 
tically 16  ft.  from  the  rail  to  the  top  of  the  second  set.  This  completes 
the  regular  raise  sets,  for  at  the  top  of  the  second  set  drifts  are  run  con- 
necting all  the  raises  in  the  section.  These  drifts  are  then  widened  from 
12  to  15  ft.,  after  which  they  are  carried  up  vertically  by  means  of  over- 
hand stoping,  the  miners  working  on  ore,  only  enough  ore  being  drawn 
off  so  a#  to  permit  them  to  be  within  easy  reach  of  the  back. 


OVERHAND    STOPING    WITH    SHRINKAGE    AND    PILLAR-CAVING 


177 


These  drifts  are  finally  carried  up  to  the  level  above,  cutting  out  a 
number  of  small  pillars  which  have  been  cut  loose  from  the  waste  above 
by  the  square-set  stope  and  are  now  partially  supported  by  the  ore 
surrounding  them. 

CONSTRUCTION  OF  CHUTES 

Formerly,  cribbed  chutes  of  8x8-in.  timbers  were  carried  up  in  the 
broken  ore  with  a  manway  compartment,  2  1/2x5  ft.,  and  a  chute, 
5x5  ft.  It  has  been  found  that  a  3-in.  plank  chute  is  practically  as  good, 
with  a  saving  of  considerable  timber.  The  inside  dimensions  of  the 


Mined  by  Square  Sets 
Timbers  removed. 


Next 
Section 


100  Level 


Pillar 

Supporting 
Roof 


After  Drifts  have| 
been  carried  up,  tL_ 

hatched  portion  bla^ 
ted  out  to  facilitate 

!  drawing  ore. 


f    This 

Pillar  j 
Mined  by; 
iCaving.  1 


Pillar 

Supporting 
Roof 


200  Level 


Next 
Section 


Drift  with  Inclined 
Raise  to  Mine  Pillar 

FIG.  87. — Vertical  section  of  stope,  Duluth  mine. 


TJie  Engineering  $  Mining  Journal 


combined  chute  and  manway  are  3  ft.  3  in.  by  6  ft.  The  chute  in  the 
clear  is  3  ft.  3  in.  square,  with  a  manway  2  ft.  6  in.  wide.  The  3-in. 
planks  are  placed  on  edge,  with  ends  beveled  at  45  deg.  The  dividing 
partition  is  a  3-in.  plank  which  fits  into  a  notch  cut  in  the  side  pieces. 
As  the  back  advances,  the  chutes  are  carried  up,  surrounded  with  ore. 

HORSE  OF  WASTE 

The  matter  of  handling  a  horse  of  waste  is  not  difficult,  as  it  can  be 
broken  and  easily  drawn  off  through  one  or  more  of  the  raises  that  are 
carried  up  from  the  level.  It  has  been  found  possible  in  mining  by  this 
system  to  place  the  raises  close  together,  thus  almost  entirely  eliminating 
the  wheelbarrow  by  shoveling  directly  into  a  chute. 


178 


MINING    WITHOUT    TIMBER 


DRAWING  THE  ORE 


The  next  operation  is  to  draw  the  ore.  This  is  accomplished  by  drill- 
ing holes  in  the  solid  ore  which  surrounds  the  second  square  set  in  each 
raise,  as  shown  by  the  hatched  portion  in  Fig.  89.  These  holes,  after 
being  blasted,  form  a  mill-hole  around  the  raise.  In  this  way  the  ore  is 
drawn  off  with  the  occasional  use  of  a  small  amount  of  powder.  The 
chute  planking  comes  out  with  the  ore.  The  short  pieces  are  usually 
unbroken,  while  perhaps  50  per  cent,  of  the  side  pieces  are  unbroken  and 
can  be  used  again.  By  this  means  all  the  ore  is  drawn  from  the  section 
and  the  small  pillars  are  left  standing. 

MINING  THE  PILLARS 

The  pillars  crush  down  and  break,  due  their  own  weight  and  a  few 
small  slips  that  usually  exist  in  this  class  of  porphyry  ore.  In  case  a 
pillar  does  not  break  down,  a  drift  is  run  on  the  level  underneath  it  and 


These  Drifts  carried  .up  to 
Level  above  cutting  out  Pillars 


The  Engineering  $  Mining  Journal 

Fio.  88. — Plan  of  stope,  -16  ft.  above  level.'Duluth  mine. 

a  raise  is  run  up  a  short  distance  into  the  bottom  of  the  pillar.  One 
side  of  this  raise  is  filled  with  holes,  the  base  of  the  pillar  is  blasted  out  and 
the  pillar  falls.  From  this  drift  a  new  set  of  inclined  raises  in  the  bottom 
16-ft.  block  of  ore  are  used  to  draw  off  the  ore  in  the  pillars.  These  raises 
are  merely  flat  sloping  floors  of  heavy  timbers,  with  head  room  blasted 
out  so  that  a  man  can  stand  up  and  bar  and  draw  the  rock  down  the  chute 
and  into  the  car.  The  chute  bottom  is  made  almost  flat,  so  that  the  ore 


OVERHAND    STOPING    WITH    SHRINKAGE    AND    PILLAR-CAVING  179 

cannot  run  down  it,  but  piles  up  at  the  bottom.  Large  boulders  are 
easily  plugged  and  blasted  at  the  mouth  of  the  chute  without  injury  to 
the  timbers.  Any  waste  can  be  sorted  before  it  is  loaded  into  the  cars 
and  need  not  be  mixed  with  the  ore.  The  small  boulders  are  broken 
with  hammers  before  being  loaded  into  the  cars. 

FURTHER  DEVELOPMENTS 

The  next  step  is  to  mine  out  the  sections  on  the  other  side  of  the 
large  supporting  pillars  A  and  A',  Fig.  88.  This  is  as  far  as  the  method 
has  been  worked  out  and  therefore  future  developments  will  be  watched 
with  great  interest.  There  are  several  courses  which  can  be  followed 
in  the  subsequent  mining.  If  the  back  and  the  pillars  supporting  it  are 
sufficiently  strong,  it  may  be  possible  to  mine  out  another  section  directly 
under  the  first,  from  the  300  level  to  the  200.  Again,  it  may  be  possible 
to  mine  the  supporting  pillar  by  caving  it,  as  in  the  mining  of  the  smaller 
pillars,  provided  that  the  waste  roof  will  stand  without  any  support. 
If,  however,  the  main  pillar  could  not  be  mined  in  this  way,  the  back 
over  the  sections  on  either  side  of  the  pillar  would  be  made  to  cave  in  and 
the  pillar  itself  would  be  mined  by  the  slicing  system.  If  this  last  were 
done,  the  remaining  ore  below  the  200  level  would  be  mined  by  slicing. 

The  method  has  been  considerably  changed  from  that  first  employed. 
Originally  the  section  was  mined  without  leaving  the  small  pillars.  It  was 
then  simple  overhand  stoping  on  ore.  The  back  then  was  usually  quite 
unsafe,  not  because  of  any  great  weight,  but  merely  due  to  large  masses  of 
ore  breaking  away  on  small  fractures,  which  are  common  in  almost  all 
kinds  of  porphyry.  After  one  of  these  stopes  caved,  burying  several 
men,  the  system  as  described  was  evolved.  Since  then  it  has  given  the 
greatest  satisfaction  and  as  now  employed  is  quite  safe,  as  the  men 
always  work  near  the  back  and  when  mining  piLars  are  well  protected. 

COST  OF  PILLAR  CAVING 

The  method  requires  practically  no  timber  and  the  greater  part  that 
is  necessary  can  be  used  again.  In  practically  every  step  in  this  method 
the  breaking  of  the  ore  is  done  with  the  least  possible  amount  of  powder 
and  labor,  while  the  ventilation  can  easily  be  kept  good  and  it  is  com- 
paratively safe.  For  its  application  it  is  absolutely  necessary  to  have 
a  strong,  solid  ore  and  a  strong  roof;  and  both  the  ore  and  the  waste 
roof  should  require  no  support  with  the  exception  of  a  few  stulls  to  hold 
up  small  slabs  and  loose  boulders.  It  is  also  necessary  that  the  ore  have 
definite  boundaries  and  be  large  enough  for  advantageous  work. 

The  body  should  be  large  so  that  it  can  be  divided  into  sections  and 
be  blocked  out  as  shown  in  Fig.  88,  preferably  extending  from  one  level 
to  the  next.  The  amount  of  waste  in  the  ore  must  always  be  small.  A 


180  MINING    WITHOUT   TIMBER 

small  amount  of  sorting  can  be  done  in  stopes  and  the  waste  drawn  off 
through  chutes,  but  the  proportion  of  ore  to  waste  must  always  be  large. 
Again,  if  the  ore  were  inclined  to  pack,  it  could  not  be  economically 
drawn  and  would  practically  have  to  be  mined  over  again. 

This  system  has  such  rigid  requirements  that  its  application  is  limited. 
With  the  exception  of  this  its  disadvantages  are  few  and  unimportant, 
while  on  the  other  hand  it  is  the  cheapest  method  of  mining  at  Cananea. 
The  cost  at  the  Duluth  mine  (which  is  new  and  presents  favorable  con- 
ditions) is  only  40  to  50  cents,  for  the  labor  and  timber  to  place  a  ton  of 
ore  in  the  chutes,  as  compared  with  75  to  85  cents  per  ton  at  the  nearby 
Elisa  mine.  The  Elisa  is  worked  by  the  system  of  "  overhand  stoping 
on  waste"  as  described  in  Example  22;  its  orebodies  are  also  irregular 
lenses  but  are  in  limestone  instead  of  porphyry. 


CHAPTER  XIV 
BACK-CAVING  INTO  CHUTES  OR  CHUTE-CAVING 

EXAMPLE  35. — HARTFORD  MINE,  NEGAUNEE,  MARQUETTE  RANGE, 

MICHIGAN 

(See  also  Example  13.) 

Caving   into    Rill    Chutes    on   Levels,  in   Sub-vertical   Wide    Vein. — 
The  Hartford  mine  (Oliver  Co.),  lying  a  short  distance  northeast  of 

Negaunee,  has  a  jaspillite  hangwall  and  a  soft,  hematite  ore.     On  the 

lower  levels  the  ore  shoot  is  100  ft.  wide  by  300  ft.  long,  and  its  method 

of  development  and  extraction  is  shown  in  Fig.  91. 

Drift  d  is  first  driven  in  the  foot  wall,  150  ft.  above  the  next  level 

above,  with  cross-cuts  b  to  b3  and  c  to  c3  turned  off  from  it  in  both  direc- 


Long.  Sec, 


Cross  Sec. 


FIG.  89. — Stoping  at  Hartford  mine. 

tions  at  50-ft.  intervals.  The  foot-wall  raises  r  are  put  up  to  the  stopes 
as  needed.  Stoping  is  started  from  a  30-ft.  raise  as  K,  the  most  advanced 
stope  being  at  the  end  of  the  lense  at  K3.  In  stope  K1  the  original  raise 
has  been  widened  by  breast  stoping  at  the  top,  from  which  it  was  cut 
down  by  underhand  benches  to  a  funnel  shape.  Next,  the  back  of  the 
stope  is  attacked  by  driving  a  raise  n  into  it  around  the  periphery,  leav- 

181 


182 


MINING    WITHOUT    TIMBER 


ing  enough  broken  ore  in  the  funnel  to  form  a  footing  for  the  tripod  of 
the  stoping  drill. 

Th's  peripheral  raise  leaves  a  hanging  core  vl  in  the  back  of  Kl,  which, 
as  the  raising  continues,  will  become  so  heavy  as  to  break  off  by  its  own 
weight  with  the  effect  shown  in  stope  K2.  In  K3  height  has  been  gained 
by  several  such  breaks.  Only  one  drill  is  worked  at  peripheral-raising 
in  one  stope,  and  as  its  two  runners  are  never  under  the  core,  they  run 
no  danger  from  its  downfall. 

Cross-cuts  6  are  driven  to  connect  the  peripheral  raises  with  a  foot- 
wall  raise  r  in  order  to  provide  an  entrance  into  a  stope  after  the  core  has 
fallen.  Stope  K3  should  not  be  holed  through  to  the  next  level  D  till 
the  ore  there,  corresponding  to  the  sill-floor  pillar  and  to  wedges  w,  wl 


Cross  Sec.  Side  El. 

FIG.  90. — Chute  at  Hartford  mine. 

and  w2  on  level  d  has  been  removed.  This  is  accomplished  by  putting 
up  a  raise  from  K3  to  D,  and  from  this  attacking  these  pillars  by  the  room- 
caving  system  of  Example  45,  and  throwing  the  broken  ore  down  into 
chute  K3. 

As  stopes  K3,  K2,  K1  and  K  get  higher  their  diameters  increase  and 
are  merged,  so  that  the  miners  from  there  on  work  the  peripheral  raise 
mostly  along  the  foot  and  hang  walls.  In  wide  veins  two  mills  should  be 
placed  abreast  as  m  and  m1;  but  these  also  will  merge  on  ascending. 
When  nearing  the  upper  level  D  the  ore  is  kept  near  the  back  so  that, 
when  holed  through,  any  debris  in  D  from  the  walls  of  the  emptied  stope 
above  will  not  descend  far.  By  withdrawing  the  ore  gradually  from 
adjoining  completed  mills  the  debris  can  be  kept  mostly  above  the  ore 
and  a  mixture  avoided. 

The  caved  ore  is  not  drawn  through  the  usual  spout-gates  into  tram 
cars,  but  falls  into  a  car  from  a  central  slot  in  the  roof  of  the  haulage 


BACK-CAVING   INTO    CHUTES    OR    CHUTE-CAVING  183 

drift.  To  arrange  this  enough  lagging  poles,  over  two  adjoining  drift 
caps  c  and  c1  (Fig.  &0),  are  omitted  to  give  space  for  two  12-in.  poles  a 
and  a1,  the  opening  between  which  is  covered  with  6-in.  poles  6,  which 
can  be  easily  taken  up  when  loading  a  tram  car  t  beneath.  Two  sets,  s 
and  s1,  are  placed  as  a  screen  over  this  slotgate  for  the  purpose  of  regulat- 
ing the  passage  of  the  broken  ore  sliding  down  from  the  stope  above,  as 
at  x  in  Fig.  89. 

With  this  screened  slot  there  is  no  chute  to  be  choked  by  the  huge 
masses  that  often  fall  off  the  core  above,  as  these  are  held  back  until  they 
can  be  broken  up.  The  core  'is  often  drilled  and  blasted,  when  there  is 
danger  of  its  breaking  off  in  chunks  too  large  to  be  easily  shattered  after 
dropping. 

This  method  requires  a  nearly  vertical  and  a  strong  hangwall,  but 
the  ore  is  best  adapted  when  somewhat  friable,  as  a  dense,  tough  ore 
would  tend  to  hang  up  and  cave  only  in  large  masses,  very  difficult  to 
break  up  below.  The  vein  must  be  wide  enough  to  permit  of  cutting  a 
core  which  is  large  enough  to  pay  for  the  raising  around  it.  This  method 
takes  no  more  timber  and  much  less  drilling  and  powder  than  the  under- 
ground milling  of  Example  12,  as  considerable  of  the  ore  is  crushed  in  the 
caving.  Ventilation  is  good,  and  wide  lenses  can  be  stoped  with  a  mini- 
mum of  development  work.  The  system  does  not  permit  of  underground 
sorting,  and  some  ore  is  consequently  lost  by  contamination  when  the 
filling  finally  falls  into  the  stope  mill. 

LABOR 

The  Hartford  mine,  at  the  time  it  was  visited,  was  producing,  exclu- 
sively by  back-caving,  1000  tons  in  two  shifts,  employing  24  air  drills  and 
255  men  above  and  below  ground.  This  force  included  20  men  on  dead 
work,  16  on  four  diamond  drills  and  12  men  on  the  stock-pile  loading. 
For  each  shift  there  was  a  general  mine  foreman  and  a  stope  boss  on  each 
of  the  four  levels.  The  development  work  of  sinking,  raising  and  drifting 
is  contracted  by  the  foot  and  stoping  by  the  ton,  one  mill  hole  being  let 
to  a  relay  of  four  men  for  each  air  drill.  Stope  contracts  were  being  let 
at  15  cents  or  16  cents  a  ton,  the  total  mining  cost  being  under  $1  and 
often  as  low  as  80  cents. 

EXAMPLE   36. — PIONEER   MINE,    ELY,    VERMILION  RANGE,  MINN. 
'(See  also  Example  20.) 

Sub-vertical  Wide  Vein:  Caving  into  Chutes  from  Sub-levels. — The 
Vermilion  is  the  most  northerly  of  the  Minnesota  ranges  and  strikes  N. 
70  deg.  E.  along  the  48th  parallel,  and,  though  the  ron  formation  extends 
here  disconnectedly  for  80  miles,  the  only  important  mines  yet  located  are 


184  MINING    WITHOUT    TIMBER 

near  the  towns  of  Ely  and  Soudan,  which  are  20  miles  apart.  The 
local  productive  zone  is  the  Soudan  iron  formation  of  Archaen  age,  which 
rests  on  the  Ely  greenstone  (a  basic  igneous  rock),  and  is  covered 
by  layers  of  intrusive  granite  and  porphyry.  The  Soudan  contains 
the  typical  iron  formation  rocks,  and  of  these  jaspilite  is  especially 
abundant. 

The  ores  include  very  hard,  specular  hematite  and  softer  reddish 
hematite,  more  or  less  hydrated;  but  neither  of  these  are  "paint"  ores 
in  the  sense  of  easily  staining  the  skin  red,  like  those  of  the  paint  ranges, 
the  Gogebic  and  the  Marquette.  The  iron  contents  vary  from  60  to  70 
per  cent.,  the  phosphorous  averages  0.06  per  cent,  and  the  silica  and 
moisture  about  5  per  cent.  each.  The  ore  bodies  lie  near  the  bottom  of 
the  Soudan  formation  and  follow  pitching  troughs  of  folded  Ely  green- 
stone. A  typical  deposit  is  that  of  the  Chandler-Pioneer  mine  at  Ely, 
which  follows  an  east-west  trough  and  is  covered  with  a  thick  layer  of 
barren  iron  formation,  except  at  the  west  end,  where  the  bare  outcrop 
first  revealed  it  to  mankind. 

The  Pioneer  mine  produces  at  full  capacity  1,000,000  tons  yearly  and 
has  two  working  shafts,  A  north  and  the  newer  B  south  of  the  ore  body. 
Of  these  A  is  vertical  and  20  1/3  ft.  by  7  ft.  inside  of  timbers,  the 
pump  compartment  being  4  ft.  and  the  two  skip  and  ore-cage  compart- 
ments each  5  ft.  long.  The  5-ton  skip  is  of  the  car  type  and  its  four 
wheels  run  between  two  7-in.  by  8-in.  wooden  guides  on  each  side  of  a 
skipway.  In  the  skipways  are  also  central  cage-guides  for  the  double- 
deck  cages,  which  are  placed  temporarily  over  the  skip  when  handling 
men.  When  idle  this  auxiliary  cage  rests  on  a  truck  running  on  a  track, 
supported  by  the  head-frame.  The  truck  can  be  moved  to  and  fro  by  an 
endless  rope  device  worked  by  a  small  .hand  windlass.  To  enable  the 
cages  to  be  slid  into  the  shaft  there  is  a  slot  through  both  decks  to  pass 
the  skip  rope. 

Shaft  A  is  lined  with  wood,  but  shaft  B  is  supported  by  a  steel  frame 
and  is  inclined  at  53  deg.  It  has  three  compartments,  a  pumpway  and 
two  skipways.  At  the  stations  the  skip  track  is  spiked  to  12-in.  square 
stringers,  but  elsewhere  to  longitudinal  plank  only. 

Tramming. — Two  30  h.  p.  Goodman  electric  locomotives  handle 
1500  tons  in  10  hours  at  shaft  B  on  one  level,  with  an  average  haul  of 
400  ft.  The  track  has  24-in.  gauge,  30-lb.  rails  and  0.5  per  cent,  grade 
toward  the  shaft.  A  locomotive  goes  out  in  one,  and  returns  in  the 
other,  of  the  two  parallel  drifts  along  the  ore  body.  One  motorman 
for  each  train  and  two  dump-men  at  the  shaft  pocket  is  the  operating 
force. 

Breaking  Ground. — The  air  is  conveyed  at  65  Ib.  pressure  through  an 
8-in.  pipe  for  11/2  miles  from  the  central  compressor  at  the  Sibley  mine. 
The  Little  Giant  3  1/8-in.  drills  are  used  with  -f  bits  made  of  octago- 


BACK-CAVING    INTO    CHUTES    OR    CHUTE-CAVING 


185 


nal   steel   and  operated  on  counter-weighed  tripods.     The  powder  is 
35  per  cent,  dynamite  and  is  fired  by  fuse  and  cap. 

In  the  diagrams  of  Fig.  91,  the  main  levels,  A  and  D,  are  spaced  100 
ft.  apart  vertically  with  two  sub-levels,  B  and  C,  equi-distant  between 
them.  On  the  floor  D  (see  plan,  Fig.  91)  drifts  are  run  to  block  out  50-ft. 
by  75-ft.  pillars,  made  large  to  prevent  a  premature  squeeze,  which 
might  close  the  haulage-ways.  On  sub-levels  B  and  C  drifts  are  run  lon- 
gitudinally near  the  foot  wall,  and  off  from  them  are  turned  parallel 
crosscuts  25  ft.  apart.  These  are  connected  by  vertical  raises  r,  which  are 
put  through  to  level  A  at  25-ft.  intervals.  Between  C  and  D  some  of 


Long.  Sec. 


Fia.  91.— Stoping  layout  at  Pioneer  mine. 

the  raises  are  inclined  in  order  to  reach  the  gates  in  a  fewer  number  of 
drifts.  Block  D  is  finally  removed  by  breaking  it  down  into  the  level 
100-ft.  below  with  the  aid  of  extra  raises. 

After  blocking  out,  the  caving  is  begun  in  slices,  and  first  the  slice 
A  (above  the  abandoned  level)  is  attached,  next  B  and  finally  C.  To  be 
safe,  the  first  cave-panel  of  level  A  must  be  kept  at  least  50  ft.  horizontally 
back  from  where  the  men  are  working  at  the  first  panel  of  sub-level  B, 
and  the  hang  wall  is  drawn  down  uniformly  from  the  farther  end  inward, 
along  the  whole  ore  body  before  attacking  the  next  panel  behind  the 
first.  The  attack  on  a  slice  is  begun  from  a  raise  by  driving  at  the  sub- 
level  B  a  drift  for  the  length  of  two  sets  in  three  directions,  one  toward 
the  hang  wall  and  two  longitudinally.  (See  Fig.  92.)  The  face  is  then 


186 


MINING    WITHOUT    TIMBER 


attacked  beyond  each  of  the  three  end  sets  (n,  g}  and  h)  of  the  drifts,  and 
large  open  spaces,  like  /  m  n  and  h  k  p,  are  blasted  out.  The  domed 
back  is  then  bored  with  a  pointed  bar  and  blasted  to  excavate  another 
shell  like  /  m'  n  and  h  kf  p}  and  the  boring  and  blasting  of  the  back  is 
continued  until  the  dome  becomes  too  large  and  dangerous  to  work 
beneath. 

If  the  dome  caves  partly  and  then  hangs  up  it  can  often  be  started 
again  by  exploding  8  to  10  sticks  of  powder,  held  below  the  cave  on  a 
long  stick.  Should  any  uncaved  wedges  be  left  above  level  B  they  can 
be  recovered  when  caving  C  by  raising  through  them  from  below.  In 
drawing  down  a  slice  while  caving  care  is  taken  to  draw  equally  from 

all  the  chutes,  so  that  the  caved  waste  above 
will  settle  equally  and  not  mix  with  the  ore. 
When  the  ore  falls  from  the  dome  in  big  masses 
it  has  to  be  blasted  to  allow  it  to  pass  the 
raises.  Much  ore  runs  into  the  raises  by  gravity, 
but  at  the  start  of  a  dome  it  has  to  be  shoveled. 
Timbering. — The  greenstone  walls  need  little 
support,  but  all  the  development  openings  in 
ore  must  be  closely  timbered  to  resist  the  heavy 
pressure  of  caving.  The  main  haulage  ways  are 
supported  by  three-quarter  sets,  having  caps 
and  posts,  8  ft.  long  by  1  ft.  to  2  ft.  diameter, 
with  the  latter  battered  2  in.  per  foot.  The 
raises  are  cribbed  closely  with  round  sticks, 
halved  at  ends  and  6  in.  to  10  in.  by  5  ft.  The 
caving  pressure  often  crumbles  the  drift  sets, 
but  seldom  before  their  usefulness  is  about  over. 

Little  timber  is  recovered  from  the  caved  ground,  and  that  saved  is 
only  good  for  cribbing. 

Application  of  Sub-level  Chute-caving  System. — Its  success  is  favored 
by  the  following  conditions:  The  deposit  should  be  large,  with  regular 
and  well-defined  boundaries  and  with  a  uniform  and  yielding  hang  wall 
on  not  too  steep  a  pitch.  The  ore  should  be  soft  enough  to  crush  under 
moderate  pressure,  and  should  be  all  of  the  same  shipping  grade,  as  no 
sorting  is  practised  underground.  The  value  per  ton  should  not  be  high, 
for  some  ore  is  unavoidably  mixed  with  the  caved  waste  and  lost. 

Where  applicable  it  is  a  cheap  system,  being  easily  ventilated,  safe, 
and  requiring  no  t  mber  and  but  little  powder  in  stoping.  Even  the 
development  openings  are  not  proportionately  numerous,  so  the  total 
timber,  drilling  and  powder  used  per  ton  of  ore  is  small  in  amount,  count- 
ing both  surface  and  underground  force.  The  output  per  shift  is  about 
four  tons  for  each  man  employed. 


FIG.  92. — Enlarging  stope  at 
Pioneer  mine. 


BACK-CAVING  INTO  CHUTES  OR  CHUTE-CAVING  187 

EXAMPLE  37. — UTAH  COPPER  MINE,  BINGHAM,  UTAH 
(See  also  Examples  3,  33,  41  and  43.) 

Irregular  Lenses  in  Porphyry.  Caving  into  Chutes  from  Sub-levels. — 
The  underground  workings  of  the  Utah  Copper  Company  are  opened  by- 
two  main  levels,  one  even  with  the  bottom  of  the  gulch  and  the  other  200  ft. 
above.  Formerly  the  caving  method  was  used  on  both  sides  of  the  creek, 
Fig.  93  but  now  underground  mining  is  confined  to  the  ground  to  the 
northeast  of  the  creek,  it  being  the  intention  to  use  steam  shovels  for 
mining  all  the  ore  on  the  other  side.  The  lowest,  or  transportation  level, 
is  cut  up  into  irregular  panels  about  75  ft.  square,  although  lately  the 
tendency  is  to  increase  the  size  of  these  panels  somewhat.  These  drifts 


FIG.  93. — Caved  surface,  Utah  Copper  mine. 

and  cross  drifts  are  driven  5x7  ft.  in  the  clear  for  single  track,  and  5x11  ft. 
for  double-track  service.  These  drifts  rarely  require  timbering  as  the 
ground  stands  well,  but  where  weak,  8x8-in.  sets  with  a  cap  5  ft.  in  the 
clear,  and  posts  7  ft.  3  in.  long,  having  a  batter  of  1/2  in.  to  the  ft.,  are 
used  for  single  track,  and  lOxlO-in.  timbers,  11  ft.  cap  and  same  length 
post  for  double  track.  In  driving  these  drifts  31/8  in.  drills  are  used. 
Back  holes  are  drilled  about  41/2  ft.  deep  and  cut  holes  about  51/2  ft. 
It  takes  from  7  to  10  holes  to  break  the  ground.  A  4-ft.  round  is  broken 
in  a  shift,  and  the  cost  is  from  $3  to  $3.50  per  foot. 

The  ore  is  hauled  in  trains  of  eight  or  nine  2  1/2-ton  bottom-dump 
cars  by  5-ton  electric  locomotives.  A  direct  current  of  500  volts  is  used, 
but  as  the  trolley  wire  is  placed  about  61/2  ft.  from  the  rails  there  is 
little  danger,  especially  as  at  chutes  and  low  places  where  a  person  might 
touch  the  trolley  wire  it  is  enclosed  in  an  upside  down  trough  (about 
6  in.  wide  and  having  4  to  6  in.  sideboards)  nailed  to  the  sprags  that 


188  MINING    WITHOUT    TIMBER 

support  the  trolley  wire.  The  wire  is  strung  from  4x4-in.  sprags  placed 
about  30  ft.  apart.  A  thirty-pound  rail  is  used.  The  main  loading 
chutes  are  fitted  with  double  rack  doors  so  as  to  facilitate  loading. 

The  main  raises  are  driven  5x7  ft.  in  hard  ground,  but  only  4x6  ft. 
in  soft  ground,  since  in  such  ground  they  widen  out  with  use  only  too 
quick.  In  driving  the  raises  a  2  1/4-in.  drill  run  by  one  man  is  used, 
and  it  takes  six  to  nine  holes  to  break  a  round  3  1/2  to  4  ft.  deep.  These 
main  raises  are  driven  on  contract  at  a  cost  of  $2  a  foot,  the  contractor 
furnishing  labor  only.  These  main  raises  are  driven  vertical  for  30  ft. 
and  then  are  turned  so  as  to  have  an  inclination  of  60  deg.  in  wet  ground 
and  50  deg.  in  dry  ground  in  the  direction  of  the  surface  slope.  This 
enables  the  raise  to  serve  more  ground,  while  the  change  in  angle  breaks 
the  fall  of  the  ore  and  prevents  its  packing  in  the  chute.  The  30  ft.  at  the 
bottom  that  is  vertical  has  proved  quite  sufficient  to  insure  an  even  feed 
at  the  mouth  of  the  chute.  In  extremely  soft  ground  the  chutes  are 
cribbed.  The  branch  raises  are  driven  approximately  5  ft.  square  and 
on  company  account,  2  1/4-in.  machines  being  used.  These  branches 
are  sometimes  advanced  on  a  slope  as  low  as  45  deg.  it  is  said.  All 
these  raises  are  driven  without  a  manway,  and  later,  when  mining  of 
the  ore  begins,  become  blind  chutes. 

Occasional  raises  are  put  up  to  serve  as  manways;  these  are  driven 
about  5x7  ft.  in  size  so  as  to  be  large  enough  for  air  pipe,  ladder  and 
timber-slide  These  ladders  are  made  with  wooden  legs  and  half-inch  round 
iron  for  rungs  so  that  they  will  not  be  broken  by  drills  that  may  occa- 
sionally slip  out  of  the  chain  when  being  hoisted  or  lowered.  All  timbers 
and  drills  are  hoisted  through  these  manways  by  hand. 

SUB-LEVEL  INTERVAL 

From  the  main  raises  sub-levels  are  driven,  and  most  of  them  are 
connected  with  the  surface  so  as  to  provide  excellent  ventilation.  These 
sub-levels  drifts  are  driven  4x7  ft.  in  the  clear,  and,  where  it  is  necessary, 
are  timbered  with  round  sets  having  caps  4  ft.  in  the  clear  and  posts 
7  ft.  long.  In  mucking  out  the  drifts,  wheelbarrows  are  used  unless  the 
run  is  greater  than  75  ft.  Then  tracks  are  put  in  and  cars  used. 

The  level  interval  was  17  ft.  at  first.  This  was  increased  to  25  ft.; 
then  to  30,  and  finally  to  33  ft.  In  a  few  places  an  interval  of  50  to  60  ft. 
was  tried,  but  with  so  high  a  back  of  ore  it  was  impossible  to  control  the 
caving  satisfactorily  so  a  great  deal  of  ore  was  lost.  From  this  exper- 
ience it  seems  that  35  ft.  is  the  limit  of  economic  caving  by  this  method 
in  the  Bingham  porphyry. 

When  it  is  desired  to  begin  caving,  all  the  raises  in  that  part  of  the 
mine  are  fitted  with  ordinary  chute  mouths,  and  at  the  same  time  a 
grizzly  made  of  lOxlO-in.  timbers,  spaced  to  give  openings  18  to  20  in. 
wide,  is  placed  over  the  top  of  the  raise  in  the  slice  below.  The  bulk- 


BACK-CAVING    INTO    CHUTES    OR    CHUTE-CAVING  189 

heads  are  necessary  not  only  to  prevent  large  boulders  from  getting  into 
the  main  chute  and  blocking  them,  but  also  in  order  that  the  chutes 
can  be  bulk-headed  in  case  capping  begins  to  run  from  above. 

After  the  raises  have  been  so  equipped,  the  tops  are  widened  until 
caving  is  almost  imminent.  Then  holes  are  drilled  until  it  is  certain 
that  the  roof  will  cave  when  they  are  blasted.  Adjacent  raises  are 
similarly  widened  out  and  blasted,  and  the  panel  caved.  This  begins  at 
the  boundary  and  progresses  away  from  it  as  the  different  chutes  run 
capping  and  have  to  be  bulk-headed.  All  these  raises  are  put  up  on  an 
angle  of  about  50  deg.  by  one  man  using  a  cross-bar  and  a  2  1/4-in.  ma- 
chine. He  stages  up  with  a  couple  of  sprags.  By  driving  the  raises  in- 
clined the  work  is  rendered  easier,  less  dangerous,  and  cheaper. 

The  bottoms  of  these  raises  are  placed  within  25  ft.  of  each  other  on 
the  sub-level  and  when  caving  begins  they  are  almost  together  at  the 
top.  As  can  be  seen,  a  pyramid  of  ground  with  a  base  25  ft.  square  is 
left  standing.  In  order  to  start  this  to  caving  a  chute  mouth  is  built 
and  a  raise  driven  in  the  block.  One  hole  is  kept  2  ft.  ahead  of  the 
other  holes  so  that  the  miner  will  know  when  he  approaches  loose 
ground.  So  long  as  this  hole  does  not  show  solid  ground  above,  the 
holes  are  blasted,  the  broken  ore  drawn  off,  and  drilling  begun  again.  But 
whenever  this  hole  shows  the  ground  to  be  crushed  and  broken,  the  sides 
of  the  enlarged  top  of  the  raise  are  drilled,  and  then  all  are  blasted  at  the 
same  time,  caving  the  raise.  This  is  called  by  the  miners  a  "general  fire 
or  general  blast."  The  drilling  of  one  hole  2  ft.  deeper  than  the  others 
insures  against  the  round's  leaving  only  a  shell  of  a  roof  to  catch  the 
miner  when  he  begins  to  pick  down  the  back,  for  it  has  been  learned  by 
experience  that  a  raise  with  a  solid  roof  2  ft.  thick  will  not  cave. 

The  "general  fire"  in  this  raise  usually  caves  the  whole  pyramid,  but 
in  order  to  get  the  ore  near  the  base  it  is  necessary  to  put  in  along  the 
drift  chute  mouths  quite  near  together,  so  as  to  draw  off  all  the  ore. 
This'  is  especially  the  case  where,  as  it  sometimes  happens,  a  drift  has  to 
be  driven  to  tap  a  pillar  that  has  caved  before  all  the  ore  could  be  drawn 
off.  The  ore  from  these  secondary  chutes  is  drawn  into  1200-lb.  cars 
running  on  a  12-lb.  track  and  trammed  to  the  nearest  branch  chute. 

Finally  before  abandoning  that  portion  of  the  mine,  branches  from 
the  branch  raises  in  the  block  below  are  driven  so  that  their  tops  are 
within  15  ft.  of  one  another.  The  tops  of  these  are  widened  out  into  a 
funnel  shape  by  means  of  water  holes,  and  the  bases  of  the  pillars  drilled 
so  that,  when  all  the  ore  that  can  be  obtained  from  the  ground  above 
has  been  drawn,  these  pillars  can  be  blasted  and  the  stope  caved.  The 
ore  then  runs  into  the  chute  in  the  block  below  and  is  drawn  off  through 
it.  Thus  the  ground  above  each  sub-level  is  caved  working  back  from 
the  boundary,  care  being  taken  not  to  undermine  any  portion  that  has 
not  yet  been  caved  on  the  sub-level  above. 


190  MINING    WITHOUT    TIMBER 

Toward  the  end  each  chute  runs  mixed  ore  and  capping,  easily  recog- 
nized by  its  reddish,  oxidized  appearance.  Drawing  continues  and  the 
capping  is  sorted  out  until  one  man  can  no  longer  run  20  to  25  small  cars 
(1200  lb.,  capacity)  in  a  shift.  Then  the  chute  is  abandoned.  Some- 
times the  chutes  get  hung  up,  but  a  stick  of  dynamite  soon  starts  them 
again.  In  case  the  hang-up  occurs  well  up  in  the  chute,  a  lighted  primer 
fastened  to  a  long  pole  is  placed  against  the  hang-up. 

DISADVANTAGES  OF  UTAH-COPPER  CAVING  METHOD 

There  are  many  drawbacks  to  the  caving  method  used  by  the  Utah 
Copper  Company.  The  ore  is  broken  mainly  in  drifts  and  raises.  By 
this  method  the  amount  of  development  on  each  sub-level  is  excessive,  for 
each  raise  must  be  met  by  a  drift  on  each  sub-level,  so  as  to  block  it  off 
whenever  capping  begins  to  run  into  a  chute.  Indeed,  a  more  expensive 
system  of  undercutting  a  block  would  be  hard  to  devise.  Again,  instead 
of  making  use  of  light  air-hammer  drills,  these  raises  are  driven  by  means 
of  heavy  piston  drills  mounted  on  a  bar.  In  blocking  out  the  ground, 
wheelbarrows  in  some  cases  are  used,  although  it  is  known  that  through 
all  the  drifts  some  ore  will  have  to  be  trammed  in  cars. 

For  successful  caving  it  is  necessary  to  have  the  surface  settle  as 
regularly  as  possible,  and  to  drop  it  over  as  wide  areas  as  feasible,  so  as 
to  avoid  mixing  the  ore  and  overburden  at  those  places  where  the  capping 
slides  down  past  the  ore.  Even  in  systems  where  this  is  done,  the  loss 
of  ore  is  considerable.  But  with  this  chute-caving  method  the  ore  and 
capping  mingle  together  throughout  the  height  of  ore  caved,  a  distance 
of  about  33  ft.;  for,  as  the  ore  is  drawn  off  through  "each  chute,  the  capping 
follows  down,  rubbing  past  the  rough  sides  of  the  ore  whose  fracture 
planes  have  been  opened  up  by  the  weight  thrown  on  the  pillars  in  the 
block.  This  mixes  some  ore  with  the  waste. 

Drawing  continues  until  the  chute  is  filled  with  a  mixture  of  ore  and 
waste  too  low  in  grade  to  pay  for  mining.  Then  this  chute  is  blocked 
off,  another  raise  is  put  up  to  cave  the  pillar  adjacent  to  this  chute,  and 
the  process  is  repeated.  .  This  mixing  of  ore  and  capping  becomes  still 
greater  as  the  caved  ore  is  drawn  down  past  the  capping  that  fills  the 
chutes  that  have  already  been  drawn.  The  percentage  of  ore  lost  can 
only  be  approximated,  but  considering  that  to  mine  the  ore,  said  to  be 
310  ft.  thick,  nine  sub-levels  are  necessary,  it  would  amount  to  at  least 
7  per  cent.,  and  where  only  three  or  four  sub-levels  are  caved  15  per  cent, 
to  20  per  cent.,  if  not  more. 

Another  disadvantage  of  the  Utah  Copper  method  is  that  it  is  not 
systematic.  A  raise  is  put  up  to  cave  a  pyramid  of  ground,  the  shape  of 
which  is  not  known.  Ore  is  pulled  from  the  chute  mouth  until,  owing  to 
mixture  with  capping,  the  grade  becomes  too  low  to  pay  for  mining. 


BACK-CAVING    INTO    CHUTES    OR    CHUTE-CAVING  191 

Then  similar  raises  are  put  up  to  cave  other  pillars  having  unknown  shapes. 
To  decide  upon  whether  all  the  ore  or  even  most  of  the  ore  has  been  ob- 
tained is  impossible.  Besides  the  Utah  Copper  Company,  only  the  neigh- 
boring Ohio  Copper  Company  is  using  this  method  of  mining. 

COST  OF  THE  SYSTEM 

At  the  time  observed,  the  Utah  Copper  company  was  mining  1500  to 
1600  tons  of  ore  per  day  by  caving  and  was  working  225  men  underground. 
As  it  was  said  that  as  much  ground  was  being  prepared  for  caving  as 
was  being  mined,  this  would  indicate  an  average  of  almost  7  tons 
to  the  man  underground.  The  average  pay  of  these  men  is  probably 
about  $2 . 65  a  day,  so  that  labor  alone  costs  38  cents  per  ton.  Assuming 
that  labor  makes  up  60  per  cent,  of  the  cost,  this  would  indicate  a  mining 
expense  of  63  cents  per  ton.  This  cost  represents  the  cost  of  caving 
the  ore  and  does  not  include  the  cost  of  blocking  the  ore  out,  which 
item  is  very  high,  owing  to  the  numerous  drifts  that  have  to  be  driven 
on  each  sub-level  and  because  of  the  large  number  of  raises.  Besides, 
it  is  known  that  when  the  ore  was  mined  by  the  older  system,  in  which 
the  caving  was  done  by  enlarging  rooms,  the  cost  of  mining  was  at  first 
$1.25  a  ton  and  later  it  was  reduced  to  $1.10.  Considering  that  con- 
siderable timber  is  necessary  in  obtaining  the  last  ore  in  each  slice,  the 
total  cost  of  caving  probably  exceeds  80  cents  per  ton. 


CHAPTER  XV 
BLOCK-CAVING  SYSTEM 

EXAMPLE  38. — PEWABIC  MINE,  MENOMINEE   RANGE,  MICH. 
(See  also  Examples  8  and  46.) 

Sub-vertical  Wide  Vein.  Blocks  Cut  Off  by  Underhand  Sloping. 
No.  Chutes. — This  mine  is  located  just  northeast  of  Iron  Mountain  and 
is  operated  by  the  Mineral  Mining  Company  of  Milwaukee,  Wis.  The 
main  orebody  is  a  deep  lense,  about  2,000  ft.  long  by  200  ft.  thick,  with 
a  soft  talc-slate  footwall  and  a  silicious-slate  hangwall.  It  dips  north- 
ward 76  deg.  to  90  deg.  and  is  overlaid  by  horizontally  bedded  Lake 
Superior  sandstone.  The  bulk  is  hard,  low-grade,  silicious  hematite, 
like  that  of  the  Traders  mine;  but  within  the  walls  is  also  a  large  shoot  of 
high-grade,  blue  hematite  resembling  Chapin  ore. 

Slicing. — The  high-grade  shoot  is  soft  and  is  stoped  by  slicing,  but 
with  only  five  sub-levels  between  the  levels  (125  ft.  vertically  apart). 
This  gives  a  distance  between  sub-levels  of  nearly  21  ft.  and  a  back  to 
cave,  above  the  room  cap,  of  over  12  ft.  Since  no  floors  are  laid  down 
before  caving,  as  in  other  mines,  it  might  be  thought  that  the  resulting 
contamination  would  be  disastrous;  but,  as  the  high-grade  shoot  lies 
entirely  within  the  vein,  it  is  not  waste  but  low-grade  ore  which  follows 
a  caved  back  onto  the  2-in.  plank  sollar,  from  which  the  shovelling  is 
done. 

Block-caving. — This  system  is  always  used  for  the  low-grade  ore, 
but  within  the  length  of  the  high-grade  shoot  the  former  ore  is  not  touched 
until  the  latter  has  first  been  removed  by  the  'slicing  of  Example  45. 
Block-caving  is  diagrammed  in  Fig.  94. 

A  block  is  250  ft.  long  on  the  vein  and  125  ft.  high  between  the  levels. 
The  block  is  laid  out  by  driving  a  wide  footwall  drift /and  four  cross-cuts 
c  to  c'"  about  80  ft.  apart  to  the  hangwall,  connected  by  drift  h.  Next 
raises  r  are  put  up  at  50-ft.  intervals  along  cross-cuts  c  and  c'"  to  within 
20  ft.  of  the  filling  above.  Cross-cuts  6  and  &'"  are  then  driven,  over  c 
and  c'",  at  the  top  of  these  raises  and  wooden  chutes  put  at  their  bot- 
toms to  allow  gravity  loading. 

From  cross-cuts  b  and  b"'  underhand  stopes  of  8-ft.  width  are  then  cut 
down  to  c  and  c'".  Simultaneously  the  block  has  been  undercut  by 
breast-stoping  from  the  cross-cuts  c  and  c'"  and  from  drift  h  until  it  is 
only  supported  by  small  round  ore  pillars  p  (dash  and  dot  in  the  plan, 

192 


BLOCK-CAVING    SYSTEM 


193 


Fig.  94),  except  alongside  c  and  c!"  where  transversal  strips  K  (dash  and 
dot)  are  left,  and  these  are  drilled,  to  be  broken  later.  As  the  ore  is 
hard  no  timber  need  be  used  in  this  development  of  a  block. 

The  250-ft.  block  is  now  free  at  the  top  and  practically  free  at  each 
end,  so  that  as  soon  as  the  supporting  pillars  p  and  K  are  blasted  put 
by  sections  the  caving  can  begin.  But  settling  is  slow,  and  a  month  may 
pass  before  the  block  has  reached  the  sill  floor  and  half  a  year  before  self- 
crushing  has  reduced  the  ore  to  first-size 
ready  for  removal. 

Drawing  of  the  caved  ore  is  then  begun  by 
allowing  it  to  run  from  the  face  of  the  seven 
drifts  d,  which,  like  the  reopened  cross-cuts  c 
and  c'",  have  been  driven  (closely  timbered) 
through  the  broken  ore  at  25-ft.  intervals. 
The  ore  falls  onto  plank  sollars,  to  be  shoveled 
into  tram  cars,  and  when  any  face  ceases  to 
run  ore,  and  shows  filling,  the  next  inward 
set  is  blasted  out  to  get  its  superincumbent 
ore,  and  the  withdrawal  is  continued  until  the 
cross-cuts  c  and  c'"  are  reached.  To  exhaust 
the  block  it  is  now  only  necessary  to  drive 
new  drifts  from  c  and  c'"  half  way  between 
the  caved  drifts  d,  and  recover  their  superin- 
cumbent ore  by  the  same  withdrawing  process 
and  finally  withdraw  the  crosscuts  themselves. 

During  this  drawing  there  is  at  each  drift 
face  a  crew  of  one  miner  and  ..four  muckers. 
The  miner  keeps  the  ground  open  by  blasting 
any  scaffolds,  and  watches  the  safety  of  the 
shovelers.  Two  muckers  load  and  the  others  shove  the  tram  car  to  drift 
/,  to  be  attached  to  the  endless-rope  haulage  system.  The  stoping  is 
only  conducted  during  the  open  season,  the  winter  being  occupied  in 
cutting  out  the  blocks.  For  an  output  of  1,800  tons  daily  500  men  and 
31  air  drills  are  required;  but,  as  a  large  portion  of  the  extraction  is  rich 
ore  (got  by  room-caving)  this  gives  no  average  output  for  block-caving 
alone.  It  is  estimated,  however,  that  the  total  cost  of  mining  the  lean 
ore  alone  by  block-caving  is  only  65  cents  to  75  cents  a  ton. 

Recently,  in  starting  a  top  block  just  under  the  sandstone  capping, 
it  failed  to  part  well  from  the  sandstone  when  dropped,  so  inclines  had  to 
be  put  down  from  the  footwall  drift  and  raises  put  up  from  their  ends  to 
recover  the  huge  masses  adhering  to  the  sandstone.  Such  mischances 
can  always  be  avoided  by  cutting  away  the  ore  from  any  adhering  side 
of  a  block  by  a  narrow  stope  before  dropping  it.  Here  the  ore  usually 
parts  easily  from  its  slate  walls,  so  the  isolating  stopes  need  only  be  cut 
is 


Cross.  Sec.' 
FIG.  94. — Stopiug  at  Pewabic  mine. 


194  MINING    WITHOUT    TIMBER 

at  the  block's  ends;  but  in  other  veins  it  might  be  necessary  to  cut  stopes 
along  either  or  both  walls.  This  system  can  be  also  applied  to  soft  ore  by 
the  use  of  some  timber  during  development. 

Block-caving  is  adapted  to  any  long-bedded  body  of  brittle  homo- 
geneous ore  with  tough  definite  walls  if  the  ore  parts  easily  from  the 
walls;  if  not,  the  bed  must  be  of  sufficient  width  to  allow  the  space  and 
expense  for  cutting  an  isolating  stope  on  the  clinging  wall.  If  a  wall  is 
not  tough  it  will  peel  off  and  mix  with  the  ore,  if  such  wall  is  a  hangwall 
or  a  highly  inclined  footwall.  If  a  wall  is  indefinite  a  contamination 
will  also  take  place,  unless  a  regular  isolating  stope  is  inserted  between 
the  block  of  good  ore  and  the  transition  zone. 

Given  a  suitable  deposit,  block-caving  takes  less  powder,  drilling  and 
timber  than  most  of  its  competitors.  It  is  safe,  provides  good  ventilation, 
and  a  great  and  easily  varied  output  from  one  level,  while  blocks  (if  not 
superimposed)  can  be  worked  at  the  same  time  on  different  levels.  In 
ores  which  are  very  hard  to  drill,  but  yet  are  brittle  enough  to  break 
up  by  block-crushing,  this  system  is  very  economical.  It  has  been 
successful  at  the  Pewabic  mine  for  seven  years,  in  producing  annually 
about  250,000  tons  of  the  leaner  ore. 

EXAMPLE  39. — MOWRY  MINE,  SANTA  CRUZ  COUNTY,  ARIZ. 

Sub-vertical  Pipe  in  Limestone;  Blocks  Cut  off  by  and  Caved  onto  Square- 
set  Floors. — The  town  lies  in  a  small  basin  bounded  on  all  sides,  save  one, 
by  low  hills  formed  of  the  same  Archaean  granite  and  schist  as  the  basin 
itself.  On  the  excepted  or  north  side,  the  bounding  hill  is  limestone  and 
it  is  in  the  contact  between  the  basin  granite  and  this  hill  that  the  ore- 
bodies  occur.  The  contact  runs  east  and  west  and  dips  about  80  deg. 
northerly,  while  the  abutting  limestone  dips  45  deg.  in  the  same 
direction. 

At  the  surface  the  ore  outcropped  as  a  series  of  disconnected  shoots, 
extending  along  the  contact  for  half  a  mile. 

Much  of  the  ore  is  spotty  and  crumbly  in  nature  and  is  a  mixture  of 
clay  and  CaCo3  with  Fe20s,  Mn02,  and  PbCo^,  with  most  of  the  silver  in 
the  last.  The  largest  chimney  is  180x65  ft.  in  section,  and  is  the  richest 
as  the  ore  appears  to  become  less  manganiferous  and  more  siliceous  as 
the  dike  is  receded  from.  As  in  Leadville,  the  lead  carbonates  occur 
hard  and  siliceous,  as  well  as  soft,  while  the  accompanying  iron  and 
manganese  oxides  increase  the  resemblance  of  the  two  camps.  Most 
of  the  ore  is  so  soft  as  to  run  freely  with  moderate  pressure  and  on  this 
fact  depends  the  success  of  the  excellent  caving  system  hereafter 
described. 

The  vertical  position  and  soft  nature  of  the  largest  shoot  permits 
the  use  of  the  caving  system  illustrated  in  Fig.  95.  Two  shafts  are  sunk, 


BLOCK-CAVING    SYSTEM 


195 


one  at  each  end  of  the  shoot,  in  the  walls,  and  levels  established  at 
150-ft.  vertical  intervals.  On  the  first  level,  these  shafts  are  then  con- 
nected by  drift  b,  through  the  shoot,  and  cross-cuts  d  turned  off  from  b 
spaced  on  25-ft.  centers.  From  this  drift  and  cross-cuts,  the  balance  of 
the  ore  on  the  sill  floor  can  be  removed.  All  the  cavity  is  timbered 
with  square  sets  of  10-in.  Oregon  fir,  5  ft.  square  horizontally  and  6  ft. 
10  in.  vertically  to  center  lines.  The  first  floor  above  the  sill  floor  is 
now  removed  and  the  same  size  sets  inserted, 
while  chutes  C  are  constructed  by  siding  up 
nearly  every  alternate  set  with  2-in.  planks. 
The  usual  wood  gates,  for  dumping  into  the 
1-ton  mine  cars  running  on  tracks  b  and  d 
are  now  inserted  below  each  chute. 

Caving  can  now  begin  by  removing  the 
roof  lagging  over  the  chute  sets,  so  that  the 
ore  will  sink  from  the  back  into  the  chutes. 
By  drawing  from  all  the  chutes  uniformly,  the 
ore  can  be  made  to  settle  vertically,  so  as  not 
to  distort  the  square  sets.  Occasional  angle 
braces  or  filled  sets  may  be  necessary.  Where 
the  back  is  strong  locally  and  bridges  a  chute, 
it  can  be  barred  down  by  miners  standing  on 
the  first  floor  below  in  a  vacant  set.  If  the 
chutes  get  clogged  with  boulders,  blasting  is 
resorted  to.  Where  the  back  is  too  hard  to 
cave,  as  happens  with  some  of  the  silicious  ore 
on  the  granite  footwall,  square  sets  are  carried 
up  above  the  first  floor,  high  enough  to  re- 
move it. 

By  this  system,  80  per  cent,  of  the  ore  is 
removed  without  timber,  powder,  or  filling, 

and  without  working  under  an  unsupported  back.  Its  only  disad- 
vantage is  that  but  one  level  can  be  stoped  at  a  time.  But  a  ower 
level  can  be  preparing,  while  the  first  level  is  being  stoped,  and  in  a 
body  of  this  size  with  plenty  of  cars,  the  output  need  be  limited  only  by 
the  hoisting  capacity  of  the  shafts.  At  present  from  100  to  125  tons 
are  hoisted  daily,  of  which  some  80  per  cent,  is  concentrated. 

As  the  mine  is  less  than  400  ft.  deep,  the  shafts  have  only  one  cage 
and  one  pump  compartment  and  the  hoists  are  single-drum.  Sullivan 
2  3/4-in.  rock  drills  are  used  for  development  and  for  the  hard  silicious 
stoping.  About  100  men  work  under  ground,  nearly  all  being  Mexicans 
except  the  bosses. 


WINES  AND  MINERALS. 


FIG.  95. — Sloping  at  Mowry 
mine. 


196  MINING    WITHOUT    TIMBER 

EXAMPLE  40. — DETROIT  COPPER  MINE,  MORENCI, 
GRAHAM  COUNTY,  ARIZ. 

(See  also  Examples  22,  29  and  30.) 

Irregular  Lenses  in  Porphyry;  Blocks  Cut  off  by  Cross-cuts  and 
Raises  and  Caved  into  Chutes. — Block-caving  slopes  are  laid  out  in  much 
the  same  way  as  the'  slicing  stopes  of  Example  46,  by  drifts  and  cross-cuts 
on  the  level  and  raises  to  the  top  of  the  ore,  but  mining  a  top  floor  with 
square-sets  is  unnecessary.  While  a  square-set  top  floor  with  timber 
mat  would  temporarily  afford  a  better  separation  between  waste  and 
ore,  it  soon  is  disarranged  and  the  timber  becomes  a  nuisance,  blocking 
the  chutes.  It  would  not  be  necessary  to  push  the  raises  to  the  top  of 
the  ore  if  the  waste  line  could  be  established  and  maintained  otherwise. 
If  the  raises  are  brought  up  to  the  waste,  it  is  a  good  idea  to  strip  them 
of  timber  before  the  actual  caving  begins.  The  operation  is  as  follows. 
At  a  certain  distance — 20  to  35  ft.  below  the  top  of  the  ore — a  working 
floor  is  started,  called  the  'sub/  or  sub-level.  The  height  of  these  sub- 
levels  depends  on  the  character  of  the  ore.  The  easier  and  more  regular 
the  ore  breaks,  the  higher  are  they  taken.  On  them  all  raises  are  con- 
nected by  drifts  and  cross-cuts,  leaving  blocks  about  20  ft.  square,  in  case 
of  5x7-ft.  raises,  25  and  30  ft.  between  centers.  These  blocks  are  sub- 
divided again  by  other  intermediate  drifts  and  cross-cuts  run  parallel  to 
the  first  series,  leaving  the  block  of  ore  standing  on  four  legs.  These  are 
weakened,  by  drilling  and  blasting  as  far  as  safety  permits,  and  again 
drilled  preparatory  to  the  final  blast,  as  are  also  the  backs  of  all  the  drifts. 
This  work  is  done  with  hammer-drills  (Waugh,  Shaw,  and  others). 
When  a  part  of  the  stope  has  been  prepared  in  this  way  the  pillar  and 
back  holes  are  loaded  and  blasted  electrically.  The  broken  ore  is  shov- 
eled into  the  chutes  from  that  part  of  the  stope  still  standing  and  drawn 
out  below  through  the  original  chutes,  and  through  others  that  are 
driven  up  to  near  the  sub-level  and  then  holed  through  into  the  broken 
mass  of  ore.  The  original  raises  are,  as  a  rule,  vertical  and  timbered. 
The  secondary  ones  can  be  inclined,  and  if  the  ground  is  solid  may  not 
require  timbering.  The  ore  is  drawn  out  until  mostly  waste  begins  to 
show  at  the  chute.  Occasional  blasting  to  break  boulders  will  usually 
keep  a  chute  going  as  long  as  needed.  In  principle,  this  method  of  stoping 
makes  it  possible  to  dispense  entirely  with  timber,  except  in  the  original 
raises,  where  it  can  be  recovered.  In  practice  stulls  will  generally  be 
found  necessary  here  and  there,  but  the  less  timber  used  the  better. 
Figs.  96  and  97  illustrate  the  method  of  working  a  block-caving  stope. 
Comparison  Between  Square-setting,  Slicing,  and  Block-caving. — Square- 
setting  can  be  done  under  the  best  conditions  prevailing  at  Morenci  for 
about  80  cents  per  ton  of  ore  extracted,  at  least  20  per  cent,  of  which 
would  be  for  the  timber  alone;  but  unfavorable  conditions,  such  as 


BLOCK-CAVING    SYSTEM 


197 


heavy  ground,  by  necessitating  reinforced  timbering,  careful  filling,  and 
small-sized  stopes,  make  the  cost  run  up  to  $2  and  more  per  ton  of  ore. 
Slicing,  as  in  Example  46,  would  cost,  under  favorable  conditions,  perhaps 
60  cents  per  ton.  While  requiring  nearly  the  same  amount  of  timber,  it 
increases  the  tonnage  mined  per  man  and  employs  a  greater  proportion  of 


Plan  of  Sub-Level,  Block-Caving  Stope 


Waste 


Leve/ 


Section  of  Block-Caving  Slope 

FIG.  96. — Stoping  layout  at  Detroit  mine. 

common  labor.  Heavy  ground  will  not  affect  its  operation  as  readily  as  it 
does  square -setting.  One  disadvantage  is  that  first-class  ore,  from  6  per 
cent,  copper  up,  used  for  direct  smelting,  can  not  be  sorted  out  easily.  An 
effort  has  been  made  to  accumulate  it  in  the  stope  and  to  run  it  out 
through  one  chute,  set  aside  temporarily  for  this  purpose,  but  the  prac- 


198 


MINING    WITHOUT    TIMBER 


tice  was  abandoned.  Leaving  the  leaner  parts  of  the  orebody,  which 
can  be  done  easily  in  square-setting,  is  also  rather  difficult  and  costly 
in  slicing,  as  it  breaks  up  the  continuity  of  the  mat  and  necessitates  new 
square-setting  below,  should  the  orebody  change  again  for  the  better. 
These  difficulties  make  it  necessary  to  mine  the  low-grade  ore  as  well  as 
the  other.  Block-caving  can  be  done  for  about  40  cents  per  ton.  It 
gives  the  greatest  tonnage  per  man  and  shift,  and  reduces  the  timber 
bill  to  almost  nothing,  but  it  is  apt  to  result  in  mining  large  amounts  of 
low-grade  ore  that  would  otherwise  not  be  mined,  and  which,  with  the 


_ 


S^ock  Caving 

Three  successive  stages  of  one  Bloc fc 
on  a  Suo-Level  preparatory  to 
shooting  it 


B/ock  /aid  out  tiy  connecting  the 
four  raises  0/7 


±J    BJock  subdivided  £>y  para //e/ drifts 
'a/se  C     Legs  weakened  ready  for  bfastinff 


FIG.  97. — Plan  of  stoping  details,  Detroit  mine. 

loss  in  concentration,  can  not  possibly  pay  expenses.  Heavy  ground 
affects  this  method  but  very  little.  Both  slicing  and  block-caving 
represent  a  large  initial  expense,  and  take  long  preparation  before  a 
stope  is  ready  for  extraction,  but  once  started  the  production  is 
more  centralized  and  a  very  large  tonnage  can  be  rapidly  mined  from 
one  stope. 

A  careful  sampling  should  follow  closely  the  opening  of  a  block  of 
ground  to  be  mined  by  the  caving  method,  as  mine  samples  as  ordinarily 
taken  prove  usually  to  be  higher  than  the  ultimate  mill  sample  covering 


BLOCK-CAVING    SYSTEM  199 

the  same  block  of  ground,  and  in  extraction,  much  of  the  profit  of  what 
should  be  won  by  cheap  mining  may  be  lost  in  taking  too  low  a  grade  of 
ore.  Sometimes  more  careful  sampling  might  result  in  abandoning  part 
of  a  stope  already  blocked  out,  as  too  low  in  grade  to  afford  a  profit, 
which  otherwise  might  have  been  mined  at  a  loss.  To  reap  the  full 
benefit  of  the  cheaper  working  methods,  improvements  in  mine  sampling 
and  concentrating  are  two  important  factors,  and  the  Morenci  companies, 
especially  the  D.  C.,  have  conducted  researches  in  both  branches  for 
several  years  past.  The  floor  plan  of  the  stope  on  the  different  levels 
should  be  as  simple  as  possible,  and  only  enough  drifts  opened  to  afford 
the  necessary  facilities  for  tramming  and  starting  the  raises.  The  up- 
keep of  the  raises  and  tracks  will  run  into  a  rather  heavy  repair  bill  in  any 
event,  and  this  should  be  kept  to  the  lowest  possible  point  consistent 
with  the  rapidity  of  working  and  handling  materials  and  ore.  Morenci 
is  somewhat  handicapped  in  reaping  the  full  benefit  of  the  cheaper 
caving  methods.  In  the  first  place,  its  orebodies  lack  the  regularity  and 
the  even  tenor  of  value  of  some  other  porphyry  ore  camps.  Then,  its 
orebodies  are  not  intact,  having  been  worked  in  their  richer  and  more 
accessible  parts  by  square-setting,  and  as  a  result,  in  places  there  are 
some  badly  cut  up  pieces  of  ground  which  now  remain  to  be  worked  by 
the  new  system.  Last,  but  not  least,  mining  in  Morenci  is  not  very 
speedy.  The  camp  is  somewhat  conservative;  hand-drilling  is  still  the 
prevailing  practice,  machine-drills  having  been  introduced  only  these  last 
few  years.  More  rapid  mining  would  certainly  reduce  the  cost,  espe- 
cially of  slicing,  and  make  possible  the  recovery  of  a  good  deal  of  timber 
that  is  now  blasted  and  lost,  at  the  same  time  that  it  would  increase  the 
tonnage  output  per  man.  The  work  is  now  frequently  handicapped  by 
the  closing  in  of  the  mat  due  to  settling  and  by  the  necessity  of  close 
stulling.  New  orebodies  that  afford  the  opportunity  of  close  pros- 
pecting by  boring  present  many  advantages,  as  in  these  instances  they 
might  be  mined  by  attacking  the  top  by  slicing,  while  raising  and  driving 
is  going  on  below,  thus  minimizing  the  time  required  for  keeping  these 
workings  open.  Systematic  boring,  too,  would  give  at  the  same  time  a 
more  reliable  sampling  than  by  any  other  method,  even  if  it  is  liable  to 
give  a  slightly  higher  average,  as  was  pointed  out  recently  by  L.  D. 
Ricketts  at  Cananea.  Boring  might  also  help  to  avoid  losses  in  other 
ways.  For  example,  at  Morenci  an  orebody  was  prepared  for  block-cav- 
ing, when  another  orebody  was  discovered  in  close  proximity.  By  the 
time  the  second  orebody  was  opened  the  ground  of  the  entire  vicinity 
was  breaking  and  settling,  owing  to  the  removal  of  the  first.  The  drifts 
could  not  be  kept  open  by  reinforcing  the  timbers,  even  with  use  of  angle 
braces  or  doubling.  The  raises  were  settling,  and  almost  constant 
timber-changing  and  easing  of  ground  had  to  be  done. 


200 


MINING    WITHOUT    TIMBER 


EXAMPLE  41. — COMMERCIAL  MINE,  BINGHAM,  UTAH 
(See  also  Examples  3,  33,  37  and  43.) 

Bedded  Lenses  in  Sloping  Limestone.  Blocks  Cut  off  by  Cross-cuts  and 
Raises  and  Caved  into  Chutes. — At  this  mine  there  are  large  lenses 
of  low-grade  silicious  ore,  200  to  500  ft.  long  and  30  to  70  ft.  thick 
with  limestone  walls,  and  for  them  the  caving  system  is  safely  and 
economically  applied  as  follows: 

No  timber  is  used,  except  for  chute  gates,  as  the  ore  will  sustain  itself 
over  the  small  excavations  opened.  The  mine  is  developed  by  an  in- 
cline, following  the  general  dip  of  the  limestone, 
and  along  it  are  started  drifts  at  lOOrft.  intervals. 
An  ore  lens  dips  with  the  limestone  and  it  is 
first  blocked  out  by  driving  and  raising  on  its 
footwall,  so  as  to  divide  it  into  50-ft.  blocks,  as 
do  sub-levels  a  and  b  and  raises  s  and  s'  in  block 
A  of  Fig.  98.  Extra  holes  are  put  in  the  sides 
and  back  of  these  drives,  to  be  blasted  later  as 
will  be  explarned.  Caving  begins  in  the  top 
strip  and  in  the  blocks  at  each  end  of  it,  the 
work  thus  proceeding  both  ways  toward  the 
center  of  the  lens.  Only  when  the  top  strip  is 
completely  caved  does  the  caving  begin  in  the 
end  blocks  of  the  next  lower  strip,  though  de- 
velopment work  may  go  on  anywhere. 

When  sub-dividing  a  50-ft.  block  below  caved  ground,  as  in  Fig.  98,  we 
have  as  a  start  the  sub-levels  a  and  6,  the  footwall  raises  s  and  s"  and 
the  raise  s',  down  as  far  as  sub-level  b.  .The  first  work  is  to  put  in  drift  d 
on  the  footwall  and  extend  raise  s'  from  sub-level  a  to  b.  Next  a  chute 
gate  is  placed  in  sub-level  a  below  each  of  the  raises  s,  s',  and  s",  so  that 
the  broken  ore  can  be  spouted  from  above  into  mine  cars  and  proceed 
to  some  chute  connected  with  the  main  level  below.  Vertical  raises,  r,  r', 
and  r",  are  then  carried  from  drift  d  to  the  hangwall  and  there  joined 
by  drift  df.  Along  the  hangwall  three  raises  are  now  put  in  (directly 
above  the  footwall  raises  s,  s',  and  s")  and  finally  drift  b'  is  opened. 
As  extra  holes  have  been  put  also,  where  necessary,  in  the  sides, 
back,  or  floor  of  these  openings  in  order  to  save  extra  set-ups  of  the  ma- 
chine drill,  the  further  excavation  is  simplified.  Holes  are  now  blasted 
along  the  sides  of  the  hangwall  drifts  and  raises  until  only  round  pillars 
(shown  dotted  at  A  in  Fig.  98)  remain  and  these  are  drilled.  The  foot-wall 
openings  are  then  treated  likewise  to  leave  only  drilled  pillars.  The  round 
hang-wall  p'llars  are  next  blasted  and  then  the  foot-wall  pillars,  and  the 
whole  blocks  should  then  break  off  along  vertical  plane  d  d'  and  fall 
on  the  footwall  if  the  following  points  have  been  considered : . 


BLOCK-CAVING    SYSTEM  201 

The  block  is  a  cube  of  which  the  upper  side  was  separated  at  the 
start;  by  the  given  work,  the  top  and  bottom  are  quite  free  and  the 
lower  side  cut  up  by  the  three  vertical  raises,  leaving  only  the  two  sides 
(that  have  been  rimmed  around  their  edges  by  the  extra  holes)  to  be 
sheared  by  the  caving.  Should  the  shearing  of  the  merely  rimmed 
sides  not  take  place,  one  or  more  intermediate  raises  must  be  put  in 
along  it,  parallel  to  raises  s  and  s".  At  the  next  block  B  of  this  strip, 
where  two  sides,  nstead  of  one,  are  caved  at  the  start,  only  one  side 
remains  to  be  separated  from  the  matrix  by  intermediate  raises  between 
s  and  t.  After  caving,  the  material  that  is  too  large  to  run  through  the 
gates  must  be  block-holed,  by  ascending  through  the  chute  gates  into 
the  cave.  After  the  completion  of  blocking  out  and  final  blast  ng  of  the 
raises,  enough  time  is  allowed  for  self-crushing  of  the  block  before 
beg  nning  to  draw  the  ore. 

EXAMPLE  42. — INSPIRATION  MINE,  GLOBE  DISTRICT,  ARIZ. 
(See  also  Examlpes  21  and  32.) 

Irregular  Lenses  in  Porphyry.  Drummond  System  Blocks  Cut  off  by 
Overhand  Sloping,  Ore-mat  under  Capping.  The  underlying  principle 
of  this  system  is  in  having  drawing-off  stopes  from  which  the,  ore  is  raked 
into  nearby  drawing-off  raises.  With  this  as  the  nucleus  from  which 
the  method  is  expanded,  Supt.  Drummond  has  devised  a  very  ingenious 
method  of  caving  the  ore  in  mass. 

On  the  "Tramming  Level"  (see  Figs.  99  and  100)  the  mine  is  cut  up 
into  a  series  of  blocks  75x200  ft.  The  drifts  d  running  across  the  ore- 
body  are  made  large  enough  for  electric  haulage,  while  the  other  drifts  e 
are  merely  subsidiary  drifts  for  connecting  the  haulage  drifts  with  one 
another  at  convenient  intervals.  So  these  drifts  are  placed  200  ft. 
apart  and  driven  only  the  ordinary  size,  as  none  are  intended  for  tram- 
ming except  one  used  as  a  main  longitudinal  entry.  Along  the  haulage 
ways  d  and  on  one  side  of  them,  raises  r  are  put  up  at  intervals  of  33  ft. 
These  later  become  the  loading  chutes  for  the  trains  and  so  are  fitted 
with  chute-mouths.  After  this  raise  has  been  carried  up  vertically  for 
12  ft.  two  branches  are  driven  to  the  " Mucking  Level"  50  ft.  above,  at 
such  an  angle  that  they  hole  into  that  level  at  such  a  distance  from  the 
mucking  openings,  which  connect  with  the  drawing-off  stopes  s,  that  the 
broken  ore  will  run  out  almost  up  to  the  edge  of  the  raise,  and  yet  is 
prevented  by  its  angle  of  repose  from  quite  getting  far  enough  to  enter 
the  chute  that  the  raise  becomes  later.  During  mining,  these  chutes 
will  be  covered  with  grizzlies  having  an  opening  14  in.  wide  between 
bars. 

On  the  Mucking  Level,  the  orebody  will  be  cut  up  into  blocks  by  a 
series  of  drifts  /  and  /',  and  cross-cuts  s.  The  drifts  /  (directly  over  the 


202 


MINING    WITHOUT    TIMBER 


f  /A»  f 


.f   to  r  ;cv\f  /ye  ±' 

v  /¥^ 


Tramming  Level 


75 


FIG.  99. — Long,  section  of  stope,  Inspiration  mine. 


Tramming  Level 
FIG.  100. — Cross-section  of  stope  (drawing  off  stopes  are  being  enlarged),  Inspiration  mine. 


BLOCK-CAVING    SYSTEM  203 

haulage  drifts  d)  will  serve  later  as  entries  to  the  mucking  places 
when  drawing  begins. 

The  other  drifts  /'  will  serve  for  tramming  the  ore  from  the  chutes 
that  are  used  in  cutting  the  stopes  s,  and  also  for  entry  to  the  raises 
and  chutes  that  serve  the  narrow  stopes  b  and  c  that  are  used  to  cut  the 
orebody  into  blocks,  so  that  the  ore  will  be  more  under  control  while  it 
is  being  caved. 

The  drawing-off  stopes  s  will  be  25  ft.  wide  at  the  bottom,  and  50  ft. 
wide  at  the  top  on  the  "Tunnel  Level,"  75  ft.  above.  These  stopes 
will  go  up  vertically  for  40  ft.,  and  then  the  sides  will  be  given  a  flare. 
In  cutting  them,  raises  will  be  carried  up  at  the  ends  and  to  give  good 
ventilation  and  for  safety  during  mining  the  stopes  will  be  joined  by 
cross-cuts  like  x. 

As  the  stopes  are  carried  up,  they  will  be  widened  in  hard  ground  so 
that  only  a  narrow  pillar  is  left  between  them,  while  in  soft  ground  the 
stopes  will  be  kept  narrow.  In  this  way  the  200-ft.  block  will  be  under- 
cut by  a  series  of  narrow  stopes  with  pillars  between  so  narrow  that  they 
will  collapse  when  the  ore  is  drawn  off. 

In  the  meantime  above  the  Tunnel  Level  the  orebody  has  been  cut 
up  by  a  serie*  of  narrow  stopes  b  and  cross  stopes  c  into  a  series  of  blocks 
75  ft.  wide  and  200  ft.  long,  that  are  centered  above  the  drawing-off 
stopes  s.  To  afford  ventilation  to  these  blocking-out  stopes  6  and  c, 
and  also  to  aid  in  the  cutting  off  of  the  orebody  at  the  top  from  the 
capping,  a  few  raises  m  are  put  up  to  surface.  These  will  pay  for  them- 
selves, as  the  supplies  can  be  lowered  through  them,  while  this  work  in 
the  upper  part  of  the  orebody  is  being  done. 

Cutting  the  Ore  off  from  the  Capping 

As  soon  as  the  blocking-out  stopes  b  and  c  have  reached  the  capping, 
the  cutt  ng  off  of  the  capping  from  the  ore  begins.  This  is  done  because 
it  is  thought  that  there  will  be  less  mixture  of  the  ore  and  waste  rock  if 
there  is  a  break  between  the  two.  Then  when  the  ore  caves,  it  does 
not  pull  the  capping  down  with  it,  but  rather  the  capping  will  follow 
down  after  and  on  top  of  the  ore.  This  cutting  of  the  ore  free  from 
the  cap  rock  is  done  by  means  of  a  series  of  parallel  drifts  k  (with  a 
pillar  12.5  ft.  wide  left  between  them)  and  cross-cuts  n.  This  pillar  is 
drilled  as  the  drifting  advances;  later,  when  the  boundary  of  the  ore 
has  been  reached,  the  pillars  will  be  blasted  retreating,  and  the  capping 
dropped.  These  drifts  will  be  driven  in  the  ore  immediately  under 
the  capping,  but  not  necessarily  all  on  the  same  level. 

When  the  ore  has  been  cut  off  from  the  capp'ng  and  the  orebody  is 
cut  up  into  the  75x200-ft.  blocks  in  the  manner  already  described,  the 
blocks  are  undercut  on  the  Tunnel  Level.  This  is  essential  in  order  to 


204  MINING    WITHOUT    TIMBER 

insure  that  all  the  ore  throughout  the  area  will  get  in  motion  when  the 
drawing  of  the  ore  begins,  for  the  drawing-off  stopes  do  not  touch  one 
another  by  about  25  ft.  This  undercutting  is  effected  by  opening  out 
a  series  of  rooms  between  which  pillars  are  left  which  are  small  enough 
so  that  they  will  collapse  when  the  ore  is  drawn  out  from  between  them, 
or  when  the  bottoms  of  them  are  blasted. 

When  all  this  has  been  accomplished  the  drawing  of  the  broken  ore 
in  the  stopes  s  begins.  This  causes  the  pillars  p  to  collapse  and  then 
the  whole  mass  of  ore  in  the  drawing-off  stopes  under  the  75x200-ft. 
block  is  set  in  motion.  This,  as  the  ore  is  drawn  out,  causes  the  pillars, 
above  to  collapse  and  come  down  and  the  whole  mass  of  ore  in  the  block 
itself  is  then  on  the  move.  Some 'of  it  will  break  coarse  and  the  rest  of 
it  small,  but  whenever  a  boulder  of  ore  gets  down  so  as  to  show  in  the 
drawing-off  holes  it  is  blasted.  Then  the  raking  of  the  ore  into  the 
drawing-off  chutes  is  resumed. 

The  lower  part  of  the  orebody  will  be  broken  up  by  its  fall,  due  to 
the  giving  way  of  the  pillars  p,  but  in  time,  as  mining  progresses  further 
ore  will  begin  to  come  down  in  large  masses.  These  will  come  down 
as  far  as  the  drawing-off  stopes,  and  then  will  temporarily  block  them. 
As  the  drawing  of  the  ore  progresses,  an  open  space  will  form  under  this 
immense  boulder  of  ground  until  there  is  quite  an  open  space  below  the 
boulder  roof  of  the  stope,  but  the  weight  of  the  ground  above  the  boulder, 
or  slab,  as  it  might  better  be  called,  will  finally  cause  the  arch  to  break, 
In  fact  the  beauty  of  the  system  is  that,  if  the  slab  is  thick,  the  weight 
of  the  unsupported  ore  will  cause  the  center  of  the  arch  to  begin  to 
dribble  away  so  as  to  approach  a  dome  or  arch  of  equilibrium,  as  is  the 
case  in  stopes  where  they  begin  to  cave.  This  crushing  action  will  cause 
the  ore  to  break  up  fine.  But,  as  the  drawing-off  stopes  are  large,  the 
arch  will  give  way  before  the  stope  is  drawn  empty  of  fine  ore. 

In  the  Drummond  method  there  will  be  great  wear  and  tear  on  the 
pillars  t  that  protect  the  mucking  places.  These  will  require  a  great 
deal  of  support  to  keep  them  open.  Possibly  the  mucking  drifts  g  will 
have  to  be  concreted,  or  else  especially  large  timber  sets  wLl  have  to 
be  used,  the  first  ones  of  which  are  protected  from  the  wear  of  the  ore 
by  an  iron  shield  made  of  steel  shapes. 

The  ore  will  be  drawn  as  regularly  as  possible  from  under  the  whole 
area  by  counting  the  carloads  coming  from  each  chute.  But  there 
may  still  be  a  great  irregularity  in  the  speed  with  which  the  capping 
settles,  causing  a  large  loss  of  ore.  To  prevent  this,  C.  T.  Rice  proposes 
that  the  block  to  be  caved  in  mass  not  only  be  cut  off  at  the  top  as  de- 
scribed already,  but  that  then  the  ore  be  undercut  and  the  block  dropped 
on  a  level  (as  q  y,  Fig.  90)  30  ft.  below  the  lowest  part  of  the  capping. 
This  would  form  a  protective  barrier,  or  mat,  to  keep  the  ore  and  the 
waste  apart  until  right  at  the  end  of  the  drawing  of  the  block. 


BLOCK-CAVING    SYSTEM  205 

If  this  were  insufficient,  another  cut-off  level  (as  q'y')  could  be  made 
below  to  provide  another  mat  that  would  float  or  ride  on  top  of  the  ore 
as  it  was  drawn  from  below.  The  purpose  of  the  breaks  that  this  under- 
cutting would  cause  are  several.  When  soft  ore  is  undercut  and  it 
begins  to  cave,  it  eats  up  into  a  dome  like  D  (Fig.  99)  having  sides 
with  a  very  steep  angle.  Consequently  it  is  very  poss  ble  that  domes 
filled  with  broken  ore  would  be  formed  during  the  drawing  of  the  ore 
from  the  caved  area  that  would  extend  up  to  the  capping  and  nibble  on 
into  that  before  the  ore  around  these  domes  got  much  into  motion. 
But  if  this  dome  were  broken,  as  would  be  the  cause  if  the  ore  were 
dropped  in  two  or  three  cut-off  levels  before  the  capping  were  reached, 
this  would  be  prevented  for  the  ore  around  one  of  these  domes  (as  F) 
would  be  subjected  to  a  shearing  action  whenever  the  dome  extended 
up  as  high  as  one  of  the  levels  on  which  the  ore  had  been  dropped. 
This  shearing  action  is  just  what  is  desired  so  as  to  get  the  ore  broken 
fine,  so  that  it  will  pull  evenly.  Where  the  orebody  is  thin,  the  mat  could 
be  made  of  waste  instead  of  ore  if  it  were  thought  advisable,  but  this 
would  be  dead  work  and  add  proportionally  to  the  cost  of  stoping. 
Finally  the  dropping  of  the  ore  in  a  block  on  several  different  levels 
above  the  Tunnel  Level  would  insure  that  the  ore  would  be  more  thor- 
oughly broken,  and  so  draw  more  evenly  from  under  the  capping. 

The  cost  of  blocking  out  the  ore  and  preparing  it  for  caving  should 
not  be  more  than  35  cents  a  ton,  even  when  a  top  mat  of  ore  is  provided 
and  the  block  is  cut  up  by  dropp  ng  it  at  vertical  intervals  of  50  ft. 
The  ore  can  be  mucked  into  the  dra wing-off  chutes,  including  the  boulders 
blasted,  for  not  over  10  cents  per  ton.  Consequently  the  cost,  including 
everything,  should  not  amount  to  much  over  75  cents  per  ton. 


CHAPTER  XVI 
SLICING  UNDER  MATS  OF  TIMBER  IN  PANELS 

EXAMPLE   43. — OLD   JORDAN   MINE,    BINGHAM,    UTAH 
(See  also  Examples  3,  33,  37  and  41) 

Irregular  Lenses  in  Limestone.  Also  Slicing  at  Low  Moor,  Va.,  and 
Square  Setting  at  Bingham. — For  this  slicing  the  square  sets  62/3  ft. 
high  by  5  ft.  square  in  the  clear  are  framed,  from  8x8-in.  Oregon  timber, 
in  a  Denver  Engineering  Works  double-ended  framer.  A  lense  is  first 
cut,  for  example,  at  the  200  level  by  a  drift  d  h  (see  Fig.  101)  and  two  two- 
compartment  raises  a  of  and  6  &'  are  put  up  from  it  to  the  top  of  the  ore- 
body  and  timbered  with  square  sets.  Then  the 
whole  top  floor  e  /  is  excavated  by  breast  stoping 
from  the  raises,  and  timbered  with  regular  square 
sets,  using  2-inch  plank  spreaders  between  the 
posts  instead  of  sills,  and  a  plank  floor.  Then 
holes  are  bored  near  the  base  of  half  the  posts 
(alternate  lines)  and  loaded  with  1/8  to  1/4  Ib.  of 
dynamite  each.  By  bunching  12  to  15  of  the 
fuses,  it  is  possible  for  a  dozen  men  to  light  the 
fuses  for  a  whole  floor  simultaneously. 

The  entire  back  falls  on  to  the  plank  floor,  as  its 
complete  caving  has  been  assured  by  blasting  holes 
in  any  solid  portion  of  the  back  at  the  same  time  as 
the  post  holes.  The  next  move  is  to  start  from  the 
raises  on  the  lower  floor  m  n  and  excavate  it  all, 
catching  up  the  floor  above  with  square  sets  during  the  advance.  When 
the  second  floor  is  finished,  it  is  only  necessary  to  drill  and  blast  its 
posts  and  bring  down  the  back  again  before  starting  another  slice. 
The  plank  roof  and  the  timbers  in  the  sinking  mass  seem  to  render  the 
back  cohesive  enough,  so  that  no  spiling  is  needed  for  driving  a  new 
slice.  The  excavation  and  caving  proceed  downward,  slice  by  slice, 
until  the  200-ft.  level  is  passed,  when  excavation  must  be  thereafter 
begun  from  raises  put  up  from  the  300-ft.  level. 

Though  slicing  takes  as  much  timber  as  regular  square  setting,  its 
advantage  lies  in  the  fact  that  safety  can  be  attained  without  the  use  of 
extra  filling,  which  may  be  expensive  to  obtain,  when  the  whole  orebody 
is  shipped.  In  irregular  ore  lenses  like  these,  slicing  may  lose  less  ore  than 

206 


FIG.  101.— Sloping  at  Old 
Jordan  mine. 


SLICING    UNDER    MATS    OF    TIMBER    IN    PANELS  207 

untimbered  caving.  The  ore-breaking  cost  more  than  in  square  setting, 
as  the  latter  is  overhand  stoping  and  allows  one  more  free  face  than  the 
breast  stoping  of  slicing,  which  is  broken  by  a  bottom  cut  off  the  solid. 
Only  one  floor  can  be  sliced  at  a  time,  but  a  large  production  can  be 
attained  by  increasing  the  number  of  raises.  In  panel-slicing,  waste  need 
not  be  moved  away  but  can  be  sorted  out  and  stored  where  broken. 

One  of  the  worst  troubles  of  slicing  is  poor  ventilation,  for  not  only  is 
there  no  open  airway  from  level  to  level,  but  the  rock  oxidation  hastened 
by  brecciation  and  timber  decay  of  the  sinking  back  creates  heat  and  gas, 
especially  when  a  slice  is  left  unworked  for  some  months.  In  one  such 
case  in  the  Old  Jordan,  the  temperature  is  high,  in  spite  of  an  artificial 
circulation  maintained  by  a  centrifugal  exhauster,  belted  to  a  vertical  air 
engine.  In  case  such  a  delayed  floor  becomes  too  heavy  to  hold  as  a 
whole,  it  may  be  divided  into  panels,  one  of  which  is  extracted  and  caved 
before  the  next  one  is  attacked. 

The  Old  Jordan  method  is  similar  to  the  one  used  at  Low  Moor,  Va.,*  in 
the  eighties  to  extract  a  vein  of  iron  ore  30  ft.  thick  and  dipping  60  deg. 
Not  only  was  the  ore  so  soft  as  to  require  close  timbering,  but  the  floor 
would  creep  if  excavations  were  kept  open  long.  The  hangwall  was 
also  weak,  being  a  band  of  broken  flint  and  clay.  Nevertheless  the  slicing 
system  overcame  these  obstacles.  The  stopes  at  Low  Moor  were  higher 
than  the  drifts,  being  12  to  15  ft.  high;  and  levels  were  4  to  6  stopes  or  60  to 
75  ft.  vertically  apart. 

The  slicing  system  was  not  in  use  by  the  Highland  Boy  or  Boston  Con. 
mines  at  Bingham,  which  used  square  setting  for  their  huge  sulphide 
lenses.  The  ore  is  very  heavy  and  goes  20  to  21  tons  to  a  set,  5  ft. 
square  by  6  2/3  ft.  high  in  the  clear,  so  that  many  sets  cannot  be  left  open 
at  once  without  disastrous  caves.  No  sorting  is  done,  as  all  the  ore 
broken  is  shipped,  so  the  filling  used  is  obtained  by  the  exploratory  cross- 
cuts. There  is  no  sharp  boundary  on  the  periphery  of  a  lens,  so  that  the 
pay-ore  limit  can  only  be  determined  by  assay.  A  chute  set  is  left  open 
at  50-  to  100-ft.  intervals  for  the  passage  of  the  ore  to  gates  in  the  adits 
below.  At  the  Boston  Con.,  the  mining  cost  is  given  at  only  $1.25  per 
ton,  but  such  a  low  figure  is  largely  due  to  the  high  specific  gravity  of  the 
ore.  The  timber  for  the  sets  comes  from  Oregon  already  framed  and  this 
effects  a  saving  in  freight  of  $4  per  M. 

EXAMPLE  44.     CUMBERLAND-ELY  MINES,  ELY,  NEV. 
(See  also  Example  4.) 

Irregu7ar  Lenses  in  Porphyry. — The  leached  capping  at  the  Cumber- 
land-Ely is  more  than  300  ft.  thick.  Not  only  the  capping  is  weak,  but 
it  is  sandy  and  runs  in  many  places.  Besides,  the  ore  is  not  strong 
enough  to  stand  without  timbers.  Owing  to  the  depth  of  capping,  under- 

*Trans  A,  I,  M,  E.,  May,  1888,  "Mining  in  Soft  Orebodies  at  Low  Moor." 


208 


MINING    WITHOUT    TIMBER 


ground  mining  had  to  be  resorted  to,  while  owing  to  the  richness  of  the 
ore,  which  averages  about  3.5  per  cent,  copper,  and  its  weak  nature  the 
slicing  system  of  mining  was  adopted. 

The  orebody  is  mined  through  the  Veteran  (Fig.  102)  shaft,  sunk 
in  the  limestone  at  some  distance  from  the  orebody.  This  shaft  has  four 
compartments  placed  side  by  side  and  is  475  ft.  deep. 

From  the  bottom  of  the  shaft  the  orebody  is  blocked  out  into  squares 
by  a  series  of  drifts  and  cross-drifts  50  ft.  apart.  These  drifts  are  driven 
7x8  ft.  in  size  as  mules  are  used  for  haulage.  They  cost  from  $5  to  $6  per 
ft.  From  the  drifts  at  each  corner  of  the  block  raises  are  driven  to  the 
capping.  These  raises,  owing  to  the  ore  being  moist,  are  driven  on  a  slope 


FIG.  102. — Caved  area,  Veteran  mine. 

of  75  deg.  so  as  to  prevent  the  ore  from  packing  in  them.  The  capping  is 
from  40  to  100  ft.  above  the  main  level,  which  is  driven  approximately  at 
the  bottom  of  the  orebody,  although  in  places  ore  extends  below  it. 

These  raises  are  timbered  with  6xl2-in.  close  cribbing  placed  flat  ways. 
This  cribbing  is  5x7  ft.  over  all  and  is  framed  with  a  regulation  half  joint. 
The  cribbing  is  divided  by  means  of  a  3-in.  plank,  fitting  into  a  gain  1  in. 
deep  in  the  wall  plates,  so  as  to  leave  a  manway  20  in.  wide  in  the  clear. 
Round  timbers  have  been  tried,  but  they  did  not  work  as  well  as  the 
framed  timbers.  These  raises  are  connected  by  means  of  sub-level  drifts 
so  as  to  leave  several  ways  of  getting  down  to  the  level.  They  are  gen- 
erally driven  on  company  account,  but  sometimes  on  a  footage  contract 
and  cost  about  $5  a  foot. 

From  these  raises  drifts  are  driven  to  the  capping  and  then  cross- 
drifts  are  turned  along  the  boundary  at  right  angles  to  the  first  drifts. 
These  meet  the  cross-drifts  from  the  next  raise  about  the  center  of  the 
block,  and  then  another  drift  is  driven  parallel  to  the  boundary  and 


SLICING    UNDER    MATS    OF    TIMBER    IN    PANELS  209 

alongside  the  other.  These  drifts  are  timbered  with  drift  sets  having 
round  posts  8  ft.  long  and  12xl2-in.  caps  with  a  2-in.  plank  7  ft.  long  nailed 
to  the  under  side  for  a  spreader  to  keep  the  posts  from  coming  in.  •  As 
the  drifts  advance  a  floor  of  2xl2-in.  planks  is  laid  and  also  a  track  for  the 
car  to  run  on.  These  sets  are  lagged  with  2x1 2-in  planks.  •  The  posts 
rest  on  the  ore.  Whenever  the  ore  extends  above  the  top  slice  it  is  mined 
by  double  deckers,  that  is  by  drift  sets  placed  on  top  of  the  lower  drift 
sets.  The  sets  are  spragged  one  from  the  other. 

After  one  drift  has  been  mined  along  the  boundary,  the  ore  alongside  is 
mined  out  by  means  of  a  series  of  parallel  drift  sets.  Sometimes  the  post 
of  one  set  is  made  to  serve  the  set  on  the  side,  but  generally  this  is  not 
done.  The  ground  is  quite  heavy  and  frequently  sprags  have  to  be  put 
up  to  re-enforce  the  caps.  After  several  parallel  cuts  have  been  mined 
and  the  timbers  in  the  first  cross-drift  have  begun  to  show  signs  of  crush- 
ing, another  layer  of  2x1 2-in.  planks  placed  crossways  to  the  first  is  laid 
so  as  to  form  a  mat  to  catch  the  finely  crushed  overburden.  The  posts 
are  then  bored  almost  through  and  loaded  with  half  a  stick  of  40  per  cent, 
dynamite.  These  holes  are  blasted  by  electricity,  25  to  40  at  a  time 
wired  in  series,  this  blasting  being  done  retreating.  It  has  been  found  that 
it  is  much  better  to  blast  these  posts  than  to  let  the  weight  of  the  over- 
burden crush  them  down,  for  when  the  posts  are  blasted,  one  is  sure  that 
all  the  overburden  has  been  dropped.  Besides,  the  mat  of  planks  comes 
down  more  evenly  and  there  is  less  danger  of  the  overburden  running 
while  mining  the  next  slice.  Thus  the  whole  slice  is  mined  out.  Some- 
times the  ore  is  mined  in  cuts  parallel  to  the  original  drift,  but  generally 
that  is  not  the  case. 

The  ore  is  trammed  in  small  cars  holding  about  12  cu.  ft.  and  dumped 
down  the  raise.  Of  course  before  blasting  the  timbers  the  tracks  are 
pulled  up.  The  ore  is  trammed  by  mule  to  the  shaft  pockets  in  trains  of 
large  cars  holding  1.15  dry  tons  of  ore  each. 

The  ore  is  mined  on  contract  at  65  cents  per  large  car  and  all  the 
men  working  from  one  raise  are  in  on  the  contract.  The  men  furnish 
their  own  powder,  fuse,  caps,  and  candles.  The  ore  is  soft  and  so  the 
holes  are  drilled  with  a  single  jack  or  in  the  soft  ground  augered.  Two 
short  rounds  are  generally  put  in  each  day;  one  blasted  at  noon,  the  other 
going  off.  Each  raise  is  provided  with  an  air  pipe  from  a  suction  blower. 

For  the  second  slice  the  men  drop  down  10  ft.  and  start  from  the  raises 
next  the  boundary  another  drift  to  the  capping.  On  this  slice  the  drift 
ssts  do  not  reach  to  the  floor  above  and  so  blocks  and  a  bridge  cap  are 
necessary.  The  bridge  caps  are  covered  with  2xl2-in.  plank  lagging. 
The  second  slice  is  mined  in  a  similar  manner  to  the  first,  and  the  floor  is 
planked  each  time  before  blasting.  The  third  slice  is  also  10  ft.  thick, 
but  the  lower  slices  are  17  ft.  thick.  In  mining  these,  double-deck  sets 
(one  drift  set  on  top  of  another)  must  be  used  to  support  the  floor  above. 

14 


210  MINING    WITHOHT   TIMBER 

The  overburden,  owing  to  its  being  moist,  is  said  to  consolidate  some- 
what after  it  is  dropped  so  that  there  is  not  much  ore  mixed  with  the  cap- 
ping when  it  becomes  necessary  to  mine  next  to  an  area  already  caved. 
In  the  panel  being  caved  there  is  always  at  least  one  open  set  between  the 
unmined  ore  and  the  sets  being  blasted.  Still  sometimes  in  the  first 
few  slices,  and  especially  where  the  capping  is  sandy,  the  overburden 
runs,  and  some  ore  is  lost.  Fortunately  at  the  Veteran  the  ore  and  cap- 
ping are  separated  by  a  very  distinct  line,  the  over  burden  being  reddish. 

The  round  posts  are  local  timber  that  comes  from  Ward  mountain  or 
Duck  creek  and  costs  about  12  to  14  cents  per  running  ft.  for  logs  9  to  12 
in.  in  diameter.  The  square  caps  come  from  Oregon.  Most  of  the  de- 
velopment work  is  done  on  company  account.  The  company  employed 
about  400  men  underground  on  the  two  shifts  and  mined  about  1500  tons 
a  day  or  3  3/4  tons  per  man. 

EXAMPLE   45. — OVERSIGHT  AND  OTHER   MINES    AT  CANANEA,  MEXICO 
(See  also  Examples  6,  18  and  34.) 

Irregular  Lenses  in  Porphyry. — Slicing  System. — Formerly  all  the  un- 
derground wide  bodies  of  ore  were  mined  by  square  sets,  but  this  method 
of  working  an  orebody  is  only  used  at  the  Kirk  mine  at  present,  where  a 
narrow  vein  with  weak  wa^s  and  a  soft  ore  re  ^uires  this  method.  It  is 
a' so  used  in  mining  the  rich  ore  capping  the  upper  part  of  the  orebodies, 
such  as  is  the  character  stic  condition  at  Cananea.  This  capping,  which 
is  generally  somewhat  strengthened  by  the  precipitation  of  new  material 
in  it,  s  often  tenacious  enough  so  that  the  timbers  can  be  robbed  before 
the  back  comes  in.  The  posts  and  caps  are  framed  from  lOxlO-in. 
Oregon  pine  and  the  girts  are  simply  8xlO-in.  pieces  sawed  to  the  right 
length.  The  posts  have  a  horn  5  in.  long  and  5  in.  square  on  each  end, 
while  the  caps  are  framed  to  have  a  5xlO-in.  horn  11/2  in.  long.  The 
sets  are  5-ft.  centers,  capways  and  girtways,  while  the  posts  are  7  ft.  4 
in.  over  all  in  the  stopes  and  framed  to  give  8  ft.  in  the  clear  on  the  sill 
floor.  The  total  cost  of  mining  with  the  square  sets  is  about  $3.25  a 
ton  where  the  timbers  are  not  recovered.  The  square-set  method  is  also 
used  in  conjunction  with  the  slicing  system  to  mine  tongues  of  ore  that 
extend  beyond  the  ore  above,  and  in  working  under  large  masses  of  waste 
that  occasionally  occur  in  the  stopes  and  are  left  unmined. 

But  the  typical  method  of  mining  at  Cananea  is  the  slicing  system. 
This  is  used  in  mining  .the  Oversight,  the  Veta  Grande  (see  Fig.  103),  the 
America  and  the  pillars  at  the  Duluth  mine,  as  well  as  the  old  gob-filled 
square-set  stopes  at  the  Capote.  The  ore  in  the  rotten,  kaol'nized  por- 
phyry at  the  Veta  Grande  and  the  Oversight  is  so  heavy  that  it  is  useless 
to  try  to  mine  with  unfilled  stopes,  while  if  the  square-set  stopes  are  filled 
this  costs  almost  as  much  as  breaking  the  soft,  rotten  ore.  Besides, 


SLICING   UNDER   MATS    OF   TIMBER   IN    PANELS  211 

owing  to  the  fact  that  the  ore  consists  of  seams  of  chalcocite  through  a 
fractured  rock  mass,  the  fines  contain  most  of  the  values,  and  with  the 
square-set  method  much  of  the  fines  are  lost  through  floors  and  go  into 
the  gob.  In  fact,  it  is  because  of  these  losses  which  attended  square-set 
mining  that  the  old  filled  square-set  stopes  in  the  Capote  and  the  Elisa 
mine  are  being  worked  again  by  slicing.  These  old  fillings  as  mined 
average  about  3  per  cent,  copper,  when  about  30  per  cent,  is  sorted 
out  in  the  stope  as  waste. 

In  mining  by  the  slicing  method  it  is  necessary  to  get  the  back 
broken  and  following  down  in  a  flat,  regular  mass  over  large  areas.  To 
accomplish  this  and  to  permit  of  getting  all  the  rich  ore  that  occurs  on 
top  of  practically  every  orebody  at  Cananea,  the  ore  above  the  first 


FIG.  103. — Caved  ground,  Veta  Grande  mine. 

level  on  which  the  ore  is  found  is  mined  by  means  of  square  sets.  In  the 
rotten  porphyry  it  is  not  difficult  to  get  this  back  to  caving,  but  in  hard 
ground  this  is  not  so  easy.  Still  this  top  weight  is  necessary  in  order  to 
have  enough  pressure  on  the  mat  to  hold  in  the  posts  when  the  •  ore  must 
be  blasted.  The  sill  floors  of  these  square-set  stopes  are  tightly  lagged 
so  as  to  allow  the  waste  mass  that  follows  down  to  be  caught  up  on  posts 
when  slicing  begins. 

The  slicing  commences  at  one  of  the  bounding  walls  of  the  orebody, 
generally  the  hanging,  and  is  carried  back  to  the  other  wall  in  slices  30 
to  75  ft.  wide  and  11  ft.  deep.  But  before  caving  begins  a  series  of  two- 
compartment  square-set  raises  are  carried  up  through  the  orebody 
approximately  30  ft.  apart.  And  as  much  as  possible  during  the  stoping 
connection  is  kept  open  with  two  raises  at  a  time.  These  raises  are 
lined  with  3-in.  planks. 


212  MINING    WITHOUT   TIMBER 

As  the  extraction  of  the  ore  progresses,  either  by  blasting  or  by 
picking,  according  to  conditions,  the  mat  of  timbers  that  is  following 
down  after  the  ore  is  caught  up  by  posts.  These  are  round  timbers  from 
Texas,  6  to  9  in.  in  diameter  and  10  ft.  long.  These  are  stood  on  5xlO-in. 
stringers  for  footboards  and  are  placed  under  the  stringers  on  the  floor 
above.  These  stringers  are  10  ft.  long  and  are  placed  5  ft.  apart  in  the 
direction  in  which  the  work  is  progressing,  for  of  course  the  stringers  are 
placed  parallel  to  the  face  that  is  being  advanced.  The  supporting  stulls 
are  often  10  ft.  apart  in  the  other  direction  for  only  those  needed  are 
used. 

As  has  been  said  the  height  of  the  slice  is  11  ft.  But  as  the.  sill 
pieces  mash  down  into  the  soft  ore  frequently  and  as  the  posts  must 
stand  on  a  5xlO-in.  sill  on  the  floor  below  they  are  made  a  foot  shorter 
than  the  slicing  height  so  as  to  allow  them  to  be  put  in  easily.  If  re- 
quired, blocking  is  put  in  between  the  post  and  stringer.  As  soon  as  the 
ore  is  mined  out,  a  floor  of  2xlO-in.  or  2xl2-in.  planks,  10  ft.  long  is 
laid  so  that  the  posts  can  be  shot  out  as  soon  as  the  pressure  becomes 
great.  These  planks  are  fitted  around  the  posts  and  the  posts  are  never 
stood  on  the  floor  as  they  might  crush  and  raise  the  floors  out  of  shape. 

One  miner  can  follow  along  with  a  second  slice  when  the  miner  ahead 
has  advanced  far  enough  so  that  each  is  beyond  the  influence  of  the 
caving  being  done  by  the  other.  This  varies  according  to  the  ground,  but 
generally  a  distance  of  50  ft.  between  the  two  is  sufficient.  If  desired  the 
following  slice  can  be  farther  to  one  side  or  it  can  be  on  the  slice  to  be 
taken  next  below.  This  is  immaterial. 

Whenever  the  pressure  on  the  posts  becomes  great  the  roof  is  dropped. 
This  is  done  by  drilling  with  an  air-driven  auger  an  inch  hole  3  or  4  in. 
into  each  post  and  loading  these  holes  with  about  a  third  of  a  stick  of 
dynamite.  These  posts  are  blasted  with  cap  and  fuse,  and  as  soon  as  the 
roof  has  become  quiet  again  the  men  come  back  to  work. 

The  ore  frequently  is  so  hard  that  it  has  to  be  drilled  by  machine 
and  in  that  case  of  course  a  piston  machine  has  to  be  used.  These  are 
2  1/2-in.  drills,  and  two  men  work  on  one  of  them.  Any  waste  that  is 
found  in  "the  ore  is  thrown  back  on  the  floor  in  a  part  where  the  roof  is 
about  to  be  caved,  while  if  a  large  mass  of  waste  is  met  with  it  is  worked 
around  and  then  by  means  of  square  sets  the  ore  below  is  worked  out 
until  finally  again  caving  is  resumed  under  the  boulder. 

In  case  the  ore  makes  out  farther  into  the  walls  than  is  the  case  on  the 
floor  above,  the  ore  is  followed  out  by  means  of  square  sets  to  the  new 
boundary,  a  floor  is  laid,  the  posts  blasted,  and  the  slicing  system 
resumed  on  the  next  slice.  In  case  the  pressure  on  the  posts  is  not  great 
enough  and  it  becomes  necessary  to  shoot  a  heavy  blast,  the  posts  are 
braced  by  means  of  2-in.  strips  nailed  to  them.  Before  the  slice  is  caved 
the  strips  are  knocked  off  for  use  in  another  part  of  the  slice. 


SLICING    UNDER    MATS    OF    TIMBER    IN    PANELS  213 

Formerly  the  floors  were  lapped  for  about  12  in.  instead  of  being 
laid  on  5xlO-in.  stringers,  but  the  use  of  stringers  only  entails  a  loss  of  a 
few  inches  of  timber  and  makes  a  floor  better  and  more  easily  caught 
up  on  the  slice  below.  The  round  timber  used  for  stulls  comes  from 
Texas  and  costs  $.80  per  10-ft.  stull,  but  this  low  price  is  obtained  by  not 
holding  the  sellers  close  to  a  fixed  size.  All  pieces  too  small  for  posts 
are  cut  in  two  on  a  rip  saw  at  the  mill,  and  the  half-poles  used  for  lagging. 
In  this  way  a  stronger  and  cheaper  lagging  than  split-lagging  is  obtained, 
while  the  cost  of  the  posts  is  considerably  less  than  formerly.  About 
three  tons  of  ore  are  mined  per  man  underground  at  Cananea  by  the 
slicing  system.  The  cost  of  labor  and  timber  for  placing  ore  in  chutes  is 
60  to  70  cents  a  ton,  making  this  system  cost  about  half  as  much  as 
square  setting,  calculated  on  the  same  basis. 

The  method  is  safe,  the  ventilation  is  fair,  sorting  can  be  done  in  the 
stope,  all  fines  that  go  through  the  floor  are  gotten  on  mining  the  next 
slice  and  not  lost,  as  in  the  square-set  method.  If  the  ore  were  saved, 
the  grade  would  be  lowered  too  much  owing  to  the  considerable  amount 
of  waste  in  the  ore.  About  30  per  cent,  is  sorted  out  in  the  stope  during 
slicing.  If  caved  this  waste  would  be  crushed  so  fine  that  it  could  not  be 
economically  sorted  on  a  picking  belt.  This  sorting  in  the  stope  is  one 
reason  for  the  mining  costs  not  being  somewhat  lower  than  they  are, 
but  they  are  much  lower  than  formerly,  even  as  it  is.  On  the  whole 
the  slicing  system  is  well  adapted  to  the  soft  ore  conditions  at  Cananea. 
Of  course  considerable  timber  is  used  in  slicing,  but  not  nearly  as  much 
as  in  the  former  square-set  method,  and  the  timber  used  is  of  cheap 
grade  as  it  does  not  have  to  be  especially  strong. 


CHAPTER  XVII 
SLICING  UNDER  ORE  WITH  BACK-CAVING  IN  ROOMS 

EXAMPLE  46. — LAKE  SUPERIOR  SOFT  ORE  LENSES  ON  GOGEBIC,  MESABI 
AND  MENOMINEE  IRON  RANGES 

(See  also  Examples  2,  7,  8,  and  38.) 

Lenses  under  Glacial  Drift  or  within  Rock  Walls.  Timber  Mats 
between  Slices. — This  system  is  applied  everywhere  on  the  Gogebic,  with 
differences  in  detail  according  to  the  mine.  Diagrams  for  East  Norrie 
practice  are  shown  in  Fig.  104.  Levels  A  and  D  are  driven  in  the  ore 
60  ft.  vertically  apart,  with  sub-levels  B  and  C  at  20-ft.  intervals. 
The  main  levels  have  two  parallel  haulageways  (M  and  M')  for 
electric  locomotives  and  are  connected  at  intervals  by  curved  cross-cuts 
D,  Df,  etc.  Next,  levels  A  and  B  are  connected  by  raises  r  with 
two  compartments  (chute  and  manway),  and  vertical  where  practical. 
At  each  sub-level  the  manways  are  covered  by  a  hinged  door  and  the 
chutes  by  a  horizontal  grating  of  rails,  set  6  in.  apart,  to  prevent  the  en- 
trance of  boulders.  At  the  chute  bottom  is  placed  a  gate  above  the 
haulageways  consisting  of  a  vertically  sliding  steel  plate  raised  by  a  lever. 
Finally  the  sub-level  "B"  is  blocked  out  by  drifts  BB  and  B'Bf  and 
cross-cuts  aa',  bb',  etc.,  though  where  the  ground  is  difficult  to  keep  open 
these  openings  need  be  put  in  only  just  before  stoping. 

All  excavations  must  be  supported,  and  round,  unbarked  timber  is 
used.  In  haulageways  the  three-quarter  sets  are  about  18  in.  in  diameter 
with  8-ft.  caps  and  posts  set  4  ft.  to  6  ft.  centers,  while  in  the  sub-levels 
there  are  7-ft.  caps  and  posts  of  8  in.  to  12  in.  in  diameter.  The  raises 
are  closely  cribbed  with  4-in.  to  6-in.  poles  and  made  3  1/2  ft.  to  7  ft.  inside 
with  a  central  cribbed  partition. 

After  sufficient  development,  stoping  can  be  begun,  the  ore  being 
removed  by  a  combination  of  drifting  and  retreating  caving.  Work  can 
begin  on  sub-level  B  when  the  ore  above  level  A  has  been  extracted 
and  the  caved  hangwall  rests  on  the  floor  of  A,  which  was  previously 
covered  with  a  floor  of  waste  timbers.  Starting  at  a',  a  room  with  sub- 
level,  three-quarter  sets  (7-ft.  posts  and  caps)  is  driven  sideways  to  wall 
at  s,  when  the  right-hand  posts  of  the  last  set  are  blasted  out  to  allow 
the  12-ft.  height  of  overhanging  ore  to  fall.  When  this  ore  has  been 
removed  by  shoveling  the  next  set  behind  is  blasted  out  and  the  with- 
drawing continued  until  the  cross-cut  aaf  is  reached.  At  the  Glenn 

214 


SLICING   UNDER   ORE    WITH   BACK-CAVING   IN    ROOMS 


215 


mine  (Mesabi)  the  side  of  every  third  room  only  is  covered  by  1-in.  boards 
before  starting  to  cave;  at  the  Fayal  (Mesabi)  mine  this  boarding  is  put 
on  each  room;  but  here,  at  the  East  Norrie,  no  side-boarding  at  all  is  used. 
The  next  operation  is  to  similarly  remove  rooms  s'  and  the  process 
is  then  continued  until  the  whole  panel  a-a'-s-f  has  been  caved.  When 
the  first  panel  is  suffic'ently  advanced  work  can  be  started  from  cross- 
cut bb',  beginning  at  room  a'.  Afterward  panels  cc' ',  dd'  and  eef  can 


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FIG.  104. — Sloping  at  East  Norrie  mine. 

be  started,  so  that  the  caved  ground  will  finally  advance  with  a  bench 
face  like  f-a-g-h-b'-c'.  As  though  overlying  ground  is  caved,  one 
•chute  top  after  another  is  covered  over  and  abandoned.  When  the 
caving  on  sub-level  B  has  retreated  25  ft.  or  so  from  the  wall  at  s, 
the  caving  on  sub-level  C  can  be  similarly  attacked  as  soon  as  the 
caving  above  will  allow  the  extreme  cross-haulage  way  D  to  be  aban- 
doned in  favor  of  D'. 

In  some  cases  it  is  easier  to  conduct  this  system  of  driving  the  rooms 
parallel  to  cross-cut  aaf  instead  of  to  drift  BE.  In  this  case  the 
rooms  would  be  started  on  both  sides  of  drifts  BE  and  E'Er  and 
would  meet  half  way  between  them,  while  the  caving  face  could  be 
kept  parallel  to  aa',  as  the  slicing  receded  from  the  end  wall  sf. 


216  MINING    WITHOUT   TIMBER 

The  distance  of  20  ft.  between  sub-levels,  at  the  East  Norrie,  is  unusu- 
ally large,  and  while  it  requires  less  development  it  makes  it  less  easy  to 
keep  the  ore  free  from  filling,  and  in  narrow  veins  or  near  horses  the  back 
may  hang  up.  At  the  Newport  Mine  (Gogebic),  D  shaft  (soft  ore)  the 
levels  are.  75  ft.  apart  with  five  sub-levels,  making  the  latter  only  12  1/2 
ft.  apart.  At  the  Newport  K  shaft  (hard  ore)  the  haulageways  and  a 
few  of  the  sub-level  drifts  are  timbered,  but  the  balance  of  the  openings 
are  unsupported  and  only  a  floor  of  timber  debris  is  laid  down  before 
caving.  Here  the  sub-levels  are  12  ft.  apart,  and  there  are  nine  between 
the  haulageways,  which  are  110  ft.  vertically  apart. 

Mesabi  Range,  Minn. — Here  underground  mining  is  the. choice  for 
deposits  too  small  to  warrant  stripping,  or  for  those  parts  of  great  bodies 
with  too  deep  a  mantle  to  allow  of  its  economical  removal.  It  can  be 
worked  equally  well  in  winter,  except  for  the  small  extra  cost  of  reloading 
the  stock  pile  for  shipment.  In  opening  the  mine  the  ore  from  develop- 
ment helps  pay  the  expense,  hence  much  less  additional  outlay  is  required 
than  for  a  stripping  system.  Lastly,  any  intermediate  bed  of  lean  ore  can 
be  left  underground.  The  chief  extra  operating  expense  over  competing 
systems  comes  from  the  use  of  much  timber  and  hand  shoveling.  It  is 
also  estimated  that  10  per  cent,  of  the  ore  is  lost  in  the  present  "  room- 
caving''  now  in  vogue. 

An  underground  system  for  this  range  must  be  adapted  to  a  soft, 
friable  bed  50  ft.  to  200  ft.  thick,  with  a  hang  wall  of  glacial  drift.  At 
first  square-setting  was  tried,  but  it  was  found  that  if  not  closely  filled  a 
stope  would  collapse.  As  practically  all  the  ore  body  is  hoisted,  such 
filling  could  only  be  obtained  from  the  surface.  To  save  this  expense 
the  slicing  system  was  adopted,  with  square-setting  as  an  auxiliary. 
This  is  now  the  only  system,  with  a  few  exceptions,  used  in  all  the  soft-ore 
mines  of  the  iron  ranges,  and  is  universal  on  the  Mesabi. 

At  Eveleth  the  main  levels  are  40  ft.  vertically  apart,  with  two  sub- 
levels  between,  and  as  the  drift  sets  have  7-ft.  posts,  a  shell  of  about  5  ft. 
of  ore  is  left  to  be  caved  in  each  slice.  The  best  practice  avoids  heavy  ore 
pillars  by  sinking  the  shafts  in  the  wall  rock;  but  the  haulage-ways  are  in 
ore,  and  for  these  heavy,  round  timber-sets  are  necessary  to  withstand  the 
roof  pressure.  Some  older  shafts  are  inclines,  sunk  in  the  foot  wall' 
under  the  sloping  side  of  an  ore  trough,  using  common  skips.  But  new 
shafts  are  often  sunk  vertically  on  account  of  cheaper  maintenance  and 
hoisting,  since  the  introduction  of  the  Kimberly  skip,  which  dumps 
automatically  in  vertical  ways.  In  the  Fayal  mine,  with  an  output  of 
1500  tons  daily,  about  150  men  are  employed  underground,  thus  giving 
10  tons  per  man,  which  average  is  decreased  to  six  or  eight  tons  per  man 
when  the  surface  force  is  included.  Owing  to  the  great  superficial  extent 
of  the  lenses,  care  is  taken  that  the  sub-level  drifts  should  not  be  vertically 
over  each  other  on  the  haulageways,  in  order  to  prevent  premature 


SLICING    UNDER    ORE    WITH    BACK-CAVING   IN    ROOMS  217 

collapse.  The  Glenn  mine  (Oliver  Co.)  near  Hibbing  presents  peculiar 
conditions,  and  has  been  most  systematically  laid  out.  As  the  ore  body 
lays  in  a  trough  and  is  but  80  ft.  thick  at  its  center,  only  one  electric- 
haulage  level  is  used,  following  the  bottom  of  the  trough  and  having 
parallel  haulageways,  A  and  A'  (Fig.  105)  50  ft.  apart.  Above  this  is 
a  level  BE'  at  a  44-ft.  vertical  interval,  with  two  sub-levels  between 
AA'  and  BB'. 

On  the  first  level  above  the  haulageway  the  cross-cuts  BB' ,  when 
they  strike  the  wall  rock,  are  continued  in  a  vertical  raise  (as  Be")  to 
the  next  sub-level  (or  level).  Then  c"c  is  driven  out  horizontally  till 
it  in  turn  strikes  the  side  of  trough  at  C  and  is  continued  in  another 


Ore 


Cross  Sec. 
FIG.  105. — Cross-section  of  stope,  Glenn  mine. 

raise  (cd)  up  to  sub-level  dd' '.  Each  of  these  raises  is  used  as  a  chute 
into  which  ore,  won  from  the  sides  of  the  trough,  can  be  dumped.  Thus 
the  highest  placed  ore  may  have  to  pass  through  several  raises  in  alter- 
nation with  cross-cuts  (as  d-d'-c-c"-B-B"-A)  before  reaching  the  electric 
trains  in  A  for  transmission  to  the  shaft. 

Chapin  Mine,  Menomince  Range,  Mich. — Since  its  opening  in  1880  the 
Chapin  has  been  one  of  the  largest  and  richest  of  the  "  Old  Range"  mines. 
The  deposit  consists  of  a  series  of  lenses  extending  6100  ft.,  canoe-shaped 
in  horizontal  section  and  conforming  with  the  enclosing  iron  slates,  which 
here  strike  N.  75°  W.  and  dip  70°  to  80°  N.  About  300  ft.  to  the  south 
the  Randville  dolomite  overlays  the  hangwall  slate  in  an  overturned  fold. 
The  lenses  pitch  30°  W.,  and  in  the  past  four  have  been  worked  while  a 
fifth  has  recently  been  found  to  the  south  of  the  others,  which  lie  end  to 
end  with  only  small  offsets.  The  largest  lense  is  2500  ft.  long  and  130  ft. 
wide  and  was  bottomed  at  about  1400  ft.  depth.  The  westernmost 
lense  was  formerly  divided  between  the  Ludington  and  Hamilton  com- 
panies, but  is  now  worked  with  the  Chapin  under  the  management  of  the 
Oliver  Co. 

Former  Sloping  Methods. — The  history  of  this  mine  well  illustrates 
the  evolution  of  underground  mining  at  Lake  Superior.  The  first  work 
was  done  by  stull  stoping,  in  which  the  hangwall  is  supported  by  stulls 


218  MINING   WITHOUT   TIMBER 

and  pillars  of  ore,  the  rest  being  stoped  overhand.  But  this  method 
proving  impractical,  because  of  the  width  of  vein  and  softness  of  ore  and 
walls,  it  was  replaced  by  modified  square-setting. 

With  the  latter  system  the  vein  was  laid  out  in  transversal  rooms  18  ft. 
apart  and  timbered  with  a  square-setting,  that  involved  the  use  of  con- 
tinuous sills  and  caps  16  in.  by  18  in.  and  18  ft.  long.  After  three  years' 
trial  the  growing  scarcity  of  huge  timbers  and  the  collapse  of  some  rooms 
compelled  the  abandonment  of  square-setting. 

Next  came  rock-filling,  which  involved  the  removal  of  the  ore  by 
breast  stoping  (beginning  at  one  level  and  raising  upward  to  the  next), 
each  slice  being  filled  with  rock  close  behind  the  extraction.  This  stoping 
resembled  that  now  in  vogue  at  the  Soudan  mine  (see  Example  20), 
except  that  here,  with  a  softer  and  wider  vein,  the  breast-stoping  and 
filling  had  to  be  carried  on  in  8-ft.  drifts  instead  of  the  whole  width  of  the 
vein  at  once.  Rock  for  filling  was  also  quarried  an  an  adjoining  sandstone 
quarry  (when  dead- work  rock  was  insufficient),  instead  of  letting  it 
quarry  itself  from  a  caving  hangwall,  as  at  the  Soudan.  An  additional 
expense  at  the  Chapin  was  the  great  amount  of  shovelling,  due  to  the 
necessity  of  packing  the  filling  tight  against  the  back  of  vein. 

Present  System. — All  these  disadvantages  of  rock-filling  has  caused  the 
substitution  for  it  at  the  Chapin  of  slicing,  as  the  latter  is  not  only  more 
flexible  but  saves  the  expense  of  digging  and  handling  filling  and  of  blast- 
ing nearly  half  of  the  ore.  The  main  haulageways  here  are  200  ft.  verti- 
cally apart  on  account  of  the  expense  of  cross-cutting  from  the  hoisting 
shaft  under  the  difficult  drainage  conditions.  Between  two  haulageways 
there  are  three  levels  50  ft.  apart,  and  between  these  latter  are  the  sub- 
levels  (12  1/2  ft.  apart),  from  which  the  rooms  are  turned  off.  Ore  on  the 
levels  and  sub-levels  is  handled  in  half -ton  cars  by  the  miners,  who 
dump  it  into  the  chutes  leading  to  the  haulageways. 

The  raises  are  about  50  ft.  apart  and  are  the  first  thing  to  be  started 
from  the  haulageways.  The  system  of  levels  and  sub-levels  drift  is  only 
completed  when  needed  for  stoping.  The  raises  are  cribbed  up  in  two 
compartments — a  continuous  chute  and  a  manway,  which  is  made  non- 
continuous  (by  transposing  it  to  the  opposite  side  of  chuteway  at  every 
sub-lateral)  in  order  to  lessen  the  danger  of  rock  falling  on  ascending  men. 

Output. — When  running  at  capacity  the  output  is  3500  tons  daily, 
with  a  total  of  900  to  950  men,  or  nearly  four  tons  per  man.  Most  of  the 
vein  is  first-class  ore  (58  per  cent,  iron),  but  bunches  of  second-class, 
which  are  made  low  grade  by  intercalcated  bands  of  jaspilite,  also  occur 
especially  as  a  transit  between  the  good  ore  and  the  wall  rock.  In  1907 
the  shipment  was  800,000  tons  and  the  timber  consumption  of  1,750,000 
bd.  ft.,  while  15,250,000  tons  was  the  total  output  to  January,  1908. 

Shafts. — At  first  inclined  shafts  were  sunk  in  the  ore,  as  many  as  10 
having  been  thus  started  in  the  main  lense  alone.  Next,  vertical  shafts 


SLICING    UNDER    ORE    WITH    BACK-CAVING   IN    ROOMS  219 

were  sunk  in  the  dolomite  hangwall  to  cut  the  ore  in  depth,  while  the 
latest  is  the  "Ludington  C"  or  Cornish  pump  shaft  put  down  vertically 
in  the  foot  wall  to  1525  ft.  depth.  The  last  is  10  ft.  by  21  ft.  inside  of  its 
steel  sets,  whose  wall  plates  are  6-in.  I-beams  set  2  ft.  to  5  ft.  apart, 
vertically,  according  to  ground.  There  are  four  compartments,  the 
shaft  being  divided  crosswise  into  three  divisions;  the  first,  10  ft.  by  6  ft. 
for  two  5-ton  Kimberley  skip  ways;  the  second,  10  ft.  by  5  ft.,  for  a  cage- 
way,  and  the  third  10  ft.  by  9  ft.,  for  the  Cornish  pump  pipes. 

Haulage. — In  the  main  haulageways  the  ore  is  usually  handled  in 
electric  trains  of  ten  3-ton  steel  cars,  on  a  track  in  the  East  Norrie,  or  24- 
in.  gauge  and  25-lb.  rails.  On  the  sub-levels  the  ore  is  trammed  by  the 
miners  in  cars,  which  usually  hold  one-half  to  one  ton,  but  occasionally 
(Fayal  mine)  are  of  2-ton  size. 

As  advantages  or  slicing  may  be  given  (1)  safety,  (2)  adjustability  to 
varying  output  and  irregular  bodies,  (3)  small  consumption  of  explosives, 
(4)  large  productive  capacity,  (5)  good  ventilation.  As  disadvantages 
may  be  mentioned  the  ore  lost  through  mixing  with  the  filling  (which  is 
around  10  per  cent.)  and  the  timber  consumption.  The  latter  is  as  large 
as  with  square-setting,  but  less  costly,  for  no  squaring  or  accurate  framing 
is  necessary.  Though  the  rooms  are  excavated  by  the  expensive  system 
of  breast  stoping,  the  large  fraction  of  the  output  that  is  caved  requires 
no  blasting. 

This  system  is  adapted  to  any  friable  ore  body  or  sufficient  width  to 
permit  free  descent  of  the  filling.  It  is  especially  applicable  to  thick,  flat 
bodies  (like  those  of  the  Mesabi)  and,  the  caving  being  all  done  in  narrow 
rooms,  it  is  probably  the  most  flexible  and  easiest  controlled  of  all  the 
caving  systems. 

Square-Setting. — This  method  is  now  only  used,  as  an  auxilary  to 
room-caving,  in  places  where  the  latter  would  be  awkward  to  apply.  On 
a  lower  level  of  the  East  Norrie  a  portion  of  the  ore  extends  only  50  ft. 
above  the  sill  floor,  as  it  is  cut  off  from  the  filling  above  by  a  thick  dike. 
The  insertions  of  square  sets  for  this  block  obviated  the  difficulty  of 
starting  a  new  cave  overhead  by  blasting. 

The  sets  here  are  of  round  timber  8  ft.  high  by  7  1/4  ft.  square  (to 
centers),  with  posts  about  18  in.  and  girts  and  ties  of  12-in.  diameter.  No 
filling  or  planking  (except  for  working  platforms)  is  used.  The  sets  are 
inserted  in  rooms,  extending  from  foot  to  hanging  of  the  ore  body  and 
from  the  level  up  to  the  dike.  Rooms  are  only  three  sets  long  and  a  pillar 
of  like  length  is  left  between  adjoining  rooms. 

After  several  rooms  are  finished  the  robbing  of  the  pillars  is  begun  by 
starting  their  excavation  from  the  hanging  side  and  inserting  square  sets. 
In  case  the  hangwall  should  cave  before  all  the  ore  is  extracted  the  men 
have  time  to  escape  and  the  balance  of  the  pillar  can  be  recovered  by 
raising  from  the  level  below.  When  finished,  the  square-set  excavation 


220  MINING    WITHOUT   TIMBER 

will  be  caved  by  blasting  out  a  central  post,  and  the  ground  below  can 
then  be  attacked  by  the  regular  room-caving  system. 

In  a  comparatively  thin  body  like  the  Glenn  mine  (see  Fig.  107) 
square  setting  is  also  a  convenient  aid.  It  is  most  applicable  near  the 
top  to  extract  the  ore  under  the  irregular  roof  efg  of  the  lense,  so  that 
room-caving  can  be  started  below  sub-level  cc'  from  a  horizontal  back. 
Here  the  sets  are  also  round  timbers  and  8  ft.  square  (to  centers)  and 
are  set  up  first  in  rooms  like  c'K,  which  measure  three  sets  back  from 
c',  three  sets  wide  on  each  side  of  cross-cut  cc'  and  as  high  as  top 
g  of  the  lense.  When  room  c'K  is-  finished  the  floor  and  the  sides 
showing  solid  ore  are  covered  with  1-in.  boards  (culls)  and  the  room 
caved  by  blasting  out  a  central  post.  Another  room,  on  the  same  cross- 
cut as  KK',  can  then  be  started  behind  the  cave  and  the  ore  won  with- 
out danger  of  contamination. 

DriLing. — In  soft  ore,  hand  augers  and  jumpers  are  used.  The  augers 
are  3  ft.  to  6  ft.  long  and  made  by  welding  a  T  handle  of  5/8-in.  round 
iron  to  an  auger,  twisted  from  3/16-in.  to  1  1/4-in.  steel  and  having  a 
double  point  1  1/4  in.  to  11/2  in.  wide.  The  jumpers  are  the  same 
length  and  of  two  kinds,  one  of  5/8-in.  round  steel,  with  1  1/2-in.  chisel 
point,  and  the  other  of  1-in.  round  steel  with  a  moil  point  4  in.  long,  set 
on  an  angle  with  the  shank,  in  order  to  bore  by  picking  into  soft  seams. 
In  case  a  hard  place  is  struck  with  a  jumper  the  whole  can  be  finished  by 
double-jacking;  and  the  latter  method  is  extensively  used  by  itself  in  the 
hard  ore  at  Newport  K  shaft  on  the  Gogebic.  Air  drills,  usually  with 
3-in.  pistons  and  60  Ib.  to  70  Ib.  of  air  pressure,  are  extensively  employed 
for  drilling  in  hard  ore  or  wall  rock. 

EXAMPLE  47.— MERCUR  MINE,  MERCUR,  UTAH 

Bedded  Lenses  in  Sloping  Limestone;  Timber  Mats  between  Slices. — 
The  Consolidated  Mercur  Gold  Mines  Company  operates  the  Mercur, 
Golden  Gate  and  Brickyard  mines,  together  with  a  cyanide  plant  of 
800  to  1000  tons  daily  capacity,  at  Mercur,  Utah.  Mercur,  a  town  of  but 
a  few  hundred  inhabitants,  is  near  the  southern  end  of  the  Oquirrh 
mountains,,  62.5  miles  by  rail  south  of  Salt  Lake  City  at  the  terminus 
of  the  Salt  Lake  &  Mercur  railroad. 

The  ore  deposits  occur  in  the  lower  portion  of  a  bed  5000  ft.  thick, 
known  as  the  Great  Blue  limestone.  It  has  been  intruded  by  several 
sheets  of  quartz  porphyry,  from  4  to  40  ft.  in  thickness,  which  in  general 
correspond  with  the  bedding  of  the  limestone  in  both  strike  and  dip. 

The  ore  deposits  underlie  the  sheets  of  porphyry  and  are  made  up  of 
both  altered  porphyry  and  altered  limestone.  Their  lines  of  greatest 
mineralization  coincide  closely  in  direction  with  a  series  of  nearly  vertical 
fissures  which  have  a  trend  toward  -the  northwast.  The  mineralizing 


SLICING    UNDER   ORE    WITH    BACK-CAVING   IN   ROOMS  221 

agents  are  believed  to  have  ascended  in  gaseous  form  through  these 
fissures,  depositing  the  ore  minerals  along  the  lower  contact  of  the  por- 
phyry and  impregnating  both  it  and  the  limestone,  forming  a  total 
thickness  of  workable  ore  varying  from  4  to  70  ft.  As  a  rule  only  a  few 
feet  of  the  porphyry  is  mineralized,  the  greater  portion  of  the  ore  being 
in  the  limestone. 

The  unaltered  porphyry  is  fine-grained,  compact  and  nearly  white  in 
color,  showing  a  lew  inconspicuous  phenocrysts  of  quartz,  feldspar  and 
biotite.  This  unaltered  rock,  however,  is  not  seen  near  the  ore  deposits. 
There  it  is  almost  black,  quite  soft,  and  contains  small  crystals  of  gypsum, 
while  upon  exposure  to  the  air  it  crumbles  and  has  the  semblance  of  a 
dried  mud.  The  mineralized  limestone  is  hard  and  cherty,  but  the  evi- 
dence points  to  the  fact  that  the  silicification  antedated  ore  deposition. 

THE  MERCUR  MINE 

The  Mercur  mine  is  at  the  southern  end  of  the  property.  All  of  the 
ore  is  now  taken  out  through  one  adit  which  has  a  single  18-in.  track, 
1200  ft.  of  which  are  equipped  with  electric  haulage,  while  on  the  re- 
maining 1900  ft.  horses  are  used. 

The  principal  vein  is  the  Mercur,  which  also  runs  through  the  other 
mines,  and  which  outcrops  plainly  on  the  hillsides  on  both  sides  of  Lewis- 
ton  canon.  It  is  dark  gray  in  color  and  is  characterized  by  fine  networks 
of  quartz  or  calcite  with  more  or  less  barite  running  through  it.  The 
vein  as  worked,  including  the  so-called  "lower"  vein  which  inmost  places 
is  mined  with  it,  varies  in  width  from  12  ft.  to  a  rnaximum  of  70  ft., 
with  an  average  width  of  from  20  to  30  ft.  The  dip  varies  from  10  or 
12  deg.  up  to  30  deg. 

In  the  narrower  portions  of  the  ore-body,  the  levels  are  driven  about 
20  ft.  apart  vertically,  and  the  ore  is  mined  in  open  stopes  which  may  or 
may^not  be  timbered  with  stulls.  From  the  drifts  inclines  are  put  up  in 
the  vein,  and  the  ore  is  broken  down  on  both  sides  of  them.  These 
stopes  have  no  fixed  dimensions,  the  width  depending  upon  the  extend 
of  the  shoot  and  also  upon  the  character  of  the  roof.  About  4  ft.  of  ore 
are  left  next  to  the  hanging-wall  to  support  the  roof,  and  stulls  are 
placed  wherever  necessary.  When  the  stope  has  attained  its  proper 
size,  the  ore  next  the  hanging  wall  is  extracted  by  the  retreating  system. 
The  roof  is  then  allowed  to  cave. 

The  above  method  is  employed  only  where  the  ore  has  a  thickness  of 
less  than  15  ft.  Most  of  the  orebodies  in  this  mine  have  a  thickness 
cons'derably  in  excess  of  this  figure,  and  these  are  mined  by  a  modifica- 
tion of  the  caving  system  of  Example  46. 

From  the  main  levels  raises  are  put  up  at  intervals  of  25  to  50  ft. 
(see  Fig.  108)  and,  so  far  as  possible,  are  kept  as  near  the  foot-wall  as  is 


222 


MINING    WITHOUT   TIMBER, 


permitted  by  the  angle  of  slope  at  which  the  ore  will  run  in  the  chutes. 
In  most  places  the  dip  of  the  orebody  is  so  low  that  a  series  of  chutes, 
offset  from  one  another,  is  necessary,  the  ore  being  drawn  from  one 
chute  and  trammed  into  another  nearer  the  hanging-wall. 

Where  the  orebody  is  wide,  as  many  as  three  drifts  approximately 
parallel  may  be  run  on  each  main  level,  chute-raises  being  carried  up 


Horizontal  Section  through  Upper*  Sublevel. 
FIG.  106. — Stoping  at  Mercur  mine. 

from  each  drift  in  order  to  reduce  to  a  minimum  the  labor  of  handling 
the  ore  in  the  stopes.  As  a  rule,  only  the  foot-wall  raises  are  used  for 
both  manway  and  ore-chute. 

From  the  raises,  sublevels  are  started  at  vertical  intervals  of  14  ft., 
the  development  on  each  sublevel  consisting  only  of  cross-cuts  driven  to 
both  the  foot-  and  hanging  walls.  The  work  of  caving  is  commenced  on 
the  upper  sub-level  and  continued  downward  in  successive  steps.  In 
starting  this  work  the  ends  of  the  cross-cuts  next  the  hanging  wall  are 
widened  out  until  two  or  more  of  the  cross-cuts  are  connected.  A  few 


SLICING   UNDER   ORE    WITH   BACK-CAVING    IN   ROOMS  223 

holes  are  then  drilled  in  the  roof  and  blasted  to  start  the  caving.  The 
broken  ore  is  then  shoveled  into  wheelbarrows  and  conveyed  to  the 
chutes.  When  waste  appears  in  the  broken  ore  the  intervening  pillars 
are  successively  sliced  in  rooms  to  the  hight  of  the  cross-cut  toward  the 
foot-wall,  meanwhile  holding  the  roof  temporarily  by  stulls.  Practica'ly 
all  the  work  is  done  by  hand,  so  heavy  timbers  are  not  needed  to  with- 
stand the  shocks  of  blasting.  As  the  working  faces  of  the  slices  recede, 
the  roof  is  allowed  to  cave,  only  maintaining  a  safe  working  place  for  the 
men  along  the  new  outline  of  the  pillar.  After  the  second  slice  has  been 
taken  out,  the  back  over  the  working  places  must  be  supported  by  three- 
quarter  sets  as  in  Example  46. 

When  the  ore  from  one  sub-level  or  slice  of  ground  has  been  extracted, 
caving  is  commenced  on  the  sub-level  immediately  below,  where  the  pro- 
cess is  carried  out  in  the  same  manner.  It  is  not  necessary  that  all  of 
the  ore  be  extracted  from  one  sub-level  before  caving  can  be  started  on 
another,  for  caving  is  carried  on  at  several  levels  in  the  mine,  but  not  in 
the  same  block  of  ore. 

RECOVERY  AND  COSTS  IN  CAVING  SYSTEM 

Where  the  sub-levels  are  14  ft.  apart  the  thickness  of  the  back  which 
remains  to  be  caved  varies  from  4  to  7  ft.,  the  lesser  thickness  having 
proved  tl  e  more  satisfactory.  It  is  questiored  by  the  mine  superin- 
tendent whether  better  results  would  not  be  obtained  if  the  sub-level  in- 
tervals were  not  reduced  to  12  ft.  By  caving  a  small  thickness  of  ore,  a 
higher  percentage  of  recovery  is  made,  and  the  operation  is  kept  under 
better  control.  It  will  be  seen  that  in  this  mine  from  50  to  70  per  cent, 
of  the  ore  is  extracted  by  room-stoping  and  slicing,  leaving  but  30  to  50 
per  cent,  to  be  obtained  by  caving  proper. 

Where  the  caving  is  begun  next  to  the  hanging-wall  it  is  sometimes 
difficult  to  obtain  all  of  the  ore  along  the  foot-wall,  particularly  where 
the  dip  of  the  orebody  is  low.  This  method,  however,  is  well  adapted 
to  the  conditions  prevailing  in  this  mine,  where  the  ore  is  firm  and  stands 
well  without  timbering  but  the  hanging  wall  is  weak.  In  case  the  caving 
was  commenced  next  the  foot-wall  the  pressure  upon  the  pillars  would 
become  almost  insupportable  by  the  time  the  caving  had  proceeded  for 
two-thirds  the  width  of  the  deposit. 

The  system  requires  little  powder  or  timber,  but  as  employed  here 
re^ui  es  a  large  amount  of  development  work  and  a  great  deal  of  shovel 
and  wheelbarrow  work.  The  ore  is  rather  soft  and  almost  all  of  the 
drilling  is  done  by  hand,  theie  being  but  seven  machine  drills  in  use  in 
the  three  mines.  Mining  is  done  largely  by  contract,  the  price  in  the 
Mercur  mine  varying  from  60  cents  to  $1.40  per  car  of  1  1/4  tons  de- 
livered at  the  electric-haulage  station,  while  in  the  Brickyard  mine  the 


224  MINING    WITHOUT   TIMBER 

costs  in  some  instances  run  as  high  as  $4  per  car.  On  the  Mercur  proper 
considerable  ore  was  taken  out  on  the  surface  through  glory  holes,  at  a 
cost  of  about  20  cents  per  car.  The  ore,  however,  was  of  low  grade. 
The  limit  on  ore  mined  is  placed  at  $3  for  oxidized  and  $4.65  for  base 
(su'phide)  ore,  though  much  ore  of  less  value  is  mined  and  milled.  Six- 
dollar  ore  is  considered  first-class,  while  $10  per  ton  is  high  grade. 

Development  is  done  largely  by  contract,  the  price  for  a  5x7-ft. 
drift  varying  from  $2  to  $4  per  ft.,  the  average  cost  being  13.50.  The 
contractors  furnish  all  their  supplies  and  deliver  their  ore  or  broken  rock 
to  the  electric-haulage  station.  With  hand  drilling  the  drifts  are  run 
at  a  rate  of  up  to  70  or  75  ft.  per  month.  Miners  receive  $2.75,  and 
machine  men  $3  for  an  eight-hour  shift,  while  contractors  average  about 
$3.50  per  day.  The  mine  is  operated  two  eight-hour  shifts  every  day  in 
the  week.  From  270  to  290  men  are  employed  underground  during  the 
24  hours  and  the  daily  output  ranges  from  700  to  800  tons  of  ore.  The 
average  figure  is,  therefore,  about  21/2  tons  of  ore  per  day  per  man. 

EXAMPLE  48. — KIMBERLEY  DIAMOND  MINES,  SOUTH  AFRICA 

Sub-vertical  Volcanic  Pipe.  No  Mats  between  Slices. — The  mineral 
deposit  here  is  a  great  vertical  volcanic  neck  or  pipe  200  yards  across. 
The  excavations  were  carried  to  a  depth  of  200  to  300  ft.,  by  open-cut, 
before  the  caving  of  sides  of  the  huge  pit  made  further  work  at  the 
bottom  so  dangerous  that  underground  mining  had  to  be  adopted. 
Vertical  shafts  were  'then  sunk  in  the  country  rock  or  "reef "  at  a  safe 
distance  from  the  open  pit  and  cross-cuts  driven  from  these  to  enter  the 
solid  diamond-bearing  or  "  blue  ground  "  below  the  surface  workings. 

Instead  of  attempting  to  withstand,  even  for  a  time,  the  pressure  of 
the  superincumbent  mass  of  broken  .reef,  the  first  system  introduced 
was  a  caving  in  the  back  and  a  filling  of  the  excavations  after  precious 
blue  ground  had  been  extracted.  When  numerous  small  tunnels  had 
been  driven  to  the  margin  of  the  mine,  that  is,  to  the  point  where  they 
reached  the  sides  of  the  crater,  the  blue  ground  was  stoped  on  both  sides 
of,  and  above,  each  tunnel  until  a  chamber  was  formed  extending  along 
the  surface  of  the  rock  (wall)  for  100  ft.  or  more,  with  an  average  width 
of  20  ft.,  and  about  20  ft.  high.  The  roof  of  the  chamber  or  gallery  was 
then  blasted  down  or  allowed  to  break  down  by  the  pressure  of  the  over- 
lying mass  of  broken  diamond-bearing  or  barren  debris. 

In  the  early  stages  of  underground  mining  there  was  an  enormous 
amount  of  diamond-bearing  ground  which  had  been  left  behind  when 
open  mining  was  discontinued,  and  which  had  been  crushed  either  by  the 
moving  sides  of  the  immense  opening  or  by  the  collapse  of  the  under- 
ground pillars  when  mined  by  the  old  system  (of  pillar  and  room).  It 
happened  frequently,  after  breaking  through  to  the  loose  ground  above, 


SLICING    UNDER   ORE    WITH   BACK-CAVING   IN   ROOMS  225 

that  clean  diamond-bearing  ground  would  run  down  as  fast  as  it  was 
removed  for  weeks  or  months  at  a  time.  The  galleries  would  at  times 
become  blocked  with  large  pieces  of  blue  ground,  which  had  to  be  blasted, 
and  then  a  further  run  of  blue  ground  would  follow.  When  the  blue- 
ground  was  worked  back  toward  the  center  of  the  crater,  large  boulders 
or  fragments  of  basalt  which  had  come  down  through  the  loose  reef 
from  the  surface  would  be  met  with.  This  system  of  working  would  be 
continued  until  reef  alone  came  down,  the  waste  or  reef  removed  being 
sent  to  the  surface  by  itself  and  piled  on  the  waste  dump.  It  formed 
only  an  inconsiderable  proportion  (1  to  4  per  cent.)  of  the  total  output. 
When  the  roof  caved  in,  the  gallery  was  nearly  full  of  blue  ground.  Only 
a  part  of  this  ground  was  removed  by  the  men  working  on  that  level,  the 
miners  prefering  to  take  it  out  on  the  next  level  below.  This  process  of 
mining  was  repeated  from  level  to  level  until  finally  there  was  no  more 
loose  ground  to  be  recovered.  The  cost  of  extracting  blue  ground  while 
loose  ground  existed  was  very  low. 

When  the  underground  work  had  reached  a  depth  of  800  ft.  or  more, 
a  new  danger  appeared.  The  huge  open  mines  are  filled  with  debris 
form  the  sides,  caused  by  the  removal  of  the  diamond-bearing  ground  by 
open  quarrying.  This  debris  was  composed  of  the  surface  red  soil, 
decomposed  basalt,  and  friable  shale,  which  extended  from  the  surface 
down  to  a  depth  of  about  300  ft.  In  addition  to  the  debris  from  the 
surrounding  rocks,  there  were  huge  masses  of  "floating  shale,"  resembling 
indurated  blue  clay  more  than  shale.  Large  heaps  of  yellow  ground  and 
tailings,  which  the  early  diggers  had  deposited  near  the  margin  of  the 
mines,  and  west-end  yellow  ground,  contributed  to  the  mud-making 
material.  The  black  shale  which  surrounds  the  mines  disintegrates 
rapidly  when  it  falls  into  them.  It  contains  a  small  percentage  of  carbon- 
aceous matter,  and  a  large  amount  of  iron  pyrites.  When  the  huge 
masses  of  shale  fell  into  the  open  mine,  they  frequently  ignited,  either 
by  friction  or,  more  probably,  by  spontaneous  combustion,  as  they  have 
been  known  to  do  on  the  dumps,  and  burned  for  months  and  even  years 
at  a  time.  These  masses  of  burned  shale  become  soft  clay  and  form  a 
part  of  the  mixture  which  fills  the  open  crater.  This  debris  moves 
down  as  the  blue  ground  is  mined  from  beneath  it,  and  becomes  mixed 
with  the  water  which  flows  into  the  open  mine  from  the  surrounding 
rock,  and  with  storm  water,  and  forms  mud.  This  overlying  mud  be- 
comes a  menace  to  the  men  working  in  the  levels  below.  Frequent 
"mud  rushes"  occurred  suddenly,  without  the  least  warning,  and  filled 
up  hundreds  of  feet  of  tunnel  in  a  few  minutes,  the  workmen  being 
sometimes  caught  in  the  moving  mass. 

It  became  evident  that  the  first  system  of  working  was  dangerous, 
the  men  sometimes  being,  when  a  mud  rush  took  place,  either  shut  in  or 
buried  in  the  mud  coming  from  the  opposite  end  of  the  mine.  It  was 

15 


226 


MINING    WITHOUT    TIMBER 


decided,  therefore,  to  work  the  mines  from  one  side  only,  and  to  have 
the  offsets  to  the  rock  connected  one  with  the  other  at  as  few  points  as 
would  be  consistent  with  the  ventilation  of  the  working  faces.  Main 
tunnels  are  driven  (about  100  ft.  apart)  across  the  crater  upon  its  longer 


FIG.  107. — Vertical  section  of  stopes,  Kimberly  mine. 

axis,  and  at  right  angles  to  these  small  tunnels  are  driven  out  every  30 
ft.  until  they  reach  the  hard  rock  an  the  south  side  of  the  mine.  These 
tunnels  are  widened,  first  along  the  rock,  until  they  connect  one  with 
another,  and  at  the  same  time  the  back  are  stopped  up  until  they  are 


SLICING   UNDER   ORE    WITH   BACK-CAVING   IN   ROOMS  227 

within  a  few  feet  of  the  loose  ground  >bove,  thus  forming  long  rooms 
or  rather  galleries,  filled  more  of  less  with  the  blue  ground,  upon  which 
the  men  stand  when  drilling  holes  in  the  backs.  The  working  levels 
were  at  first  30  ft.  apart  vertically,  but  for  greater  economy  the  distance 
was  soon  changed  to  40  ft. 

The  broken  blue  ground  in  the  galleries  is  taken  out,  as  a  rule,  before 
there  are  any  signs  of  the  roof  giving  away.  At  times  this  is  impossible, 
and  the  roofs  cave  upon  the  broken  ground,  and  the  blue  ground  is 
covered  with  (barren)  reef.  Fig.  107  shows  the  arrangement  of  cross-cuts 
and  drifts  and  the  manner  of  stoping  under  the  conditions  described. 

As  the  roof  caves  or  is  blasted  down,  the  blue  ground  is  removed, 
and  the  loose  reef  lying  above  it  comes  down  and  fills  the  gallery.  Tun- 
nels are  often  driven  through  the  loose  reef,  and  the  blue  ground  which 
has  been  cut  off  and  buried  by  debris  is  taken  out;  but  it  is  sometimes 
left  for  those  working  the  next  level  below  to  extract. 

After  the  first  "  cut "  near  the  rock  is  worked  out,  another  cut  is  made 
and  in  this  manner  the  various  levels  are  worked  back,  the  upper  level  in 
advance  of  the  one  below,  forming  terraces  as  shown  in  Fig.  107.  The 
galleries  are  not  supported  in  any  way  with  timbers,  but  all  tunnels  in 
soft  blue  ground  are  timbered  with  sets  of  two  props  and  a  cap  of  round 
timber,  and  are  covered  with  inch  and  a  half  lagging.  Soft  blue  ground 
is  drilled  with  junper-drills  sharpened  at  both  ends.  In  hard  blue  ground 
drills  and  single-hand  hammers,  are  used.  The  native  workers  become 
very  skilful  in  both  methods  of  drilling,  and  do  quite  as  much  work  as 
white  men  would  do  under  similar  conditions. 


CHAPTER  XVIII 
PRINCIPLES  OF  MINING  SEAMS 

(a)  COMPARISON  OF  LONGWALL  AND  PILLAR  SYSTEMS 

In  the  mining  of  seams  of  coal  or  of  other  bedded  deposits  of  similar 
regular  thickness  over  large  areas  like  iron  ore,  gypsum,  salt,  etc.,  the 
two  systems  largely  used  are  "longwall"  and  "room  and  pillar"  or 
simply  "pillar."  The  longwall  system  extracts  all  the  coal  in  the  first 
operation  along  a  long  stretch  of  "  wall "  or  face  and  allows  the  roof  to 
settle  gradually  behind  the  miners  upon  a  "  gob  "  or  partial  waste  filling. 
The  only  excavation  kept  open  besides  a  narrow  passage  at  the  working 
face  are  a  few  roads,  to  the  bottom  of  the  shaft  or  other  exit,  for  venti- 
lation and  for  the  tramming  of  coal  and  supplies.  In  the  pillar  system, 
on  the  contrary,  the  first  operation  consists  of  driving  roadways  to  ex- 
tract only  part  of  the  seam  in  rooms,  while  leaving  the  balance  in  the 
form  of  pillars  to  sustain  the  roof.  Later  the  pillars  are  drawn  or 
" robbed"  so  as  to  finally  recover  as  much  of  the  seam  as  possible. 

Either  system  may  be  pursued  "advancing"  or  "  retreating."  If 
advancing,  the  attack  of  the  longwall  or  the  robbing  of  the  pillar  system 
begins  next  the  safety  pillar,  left  to  protect  the  shaft  or  other  entrance, 
and  advances  outwardly  toward  the  boundry;  while  if  retreating,  the 
roadways  of  the  longwall  or  the  roads  and  rooms  of  the  pillar  system 
are  driven  to  the  outer  boundary  of  the  mine  before  the  "attack  of  face" 
or  the  "  robbing  of  pillars,  "  respectively,  is  begun. 

The  two  broad  divisions  of  the  longwall  system  are  "continuous- 
face,"  in  which  the  face  is  kept  in  the  form  of  a  circle  or  similar  closed 
figure,  and  "  panel "  where  the  face  is  handled  in  panels  or  blocks,  along 
a  sufficient  stretch  for  free  roof  subsidence,  without  forming  a  closed 
figure.  These  divisions  have  each  several  varieties  and  often  shade  into 
each  other. 

The  pillar  system  has  three  varieties:  "room  and  pillar,"  where  the 
rooms  are  wider  than  the  pillars;  "stall  and  pillar,"  where  the  stall  or 
room  is  narrower  than  the  pillar;  and  "panel,"  where  the  mine  is  divided 
into  sections  or  panels,  separated  from  each  other  by  peripheral  pillars, 
and  each  is  divided  into  a  number  of  rooms  with  corresponding  pillars. 
In  some  mines  it  has  been  found  advantageous  to  combine  the  longwall 
and  pillar  systems  or  even  to  operate  them  separately  in  different  portions 
of  the  property. 

The  longwall  system  is  adapted  only  to  uniform  seams  with  roofs  of 

228 


PRINCIPLES  OF  MINING  SEAMS  229 

an  elastic  material  like  shale  or  sandstone  rather  than  those  with  a  blocky 
fracture  like  limestone.  Hitherto,  longwall  has  been  most  used  for 
working  seams  of  coal,  but  it  is  likely  hereafter  to  be  widely  applied  to 
other  deposits  like  the  Appalachian  iron  beds,  the  Michigan  copper 
amygdaloids,  or  the  Transvaal  gold  banket,  in  order  to  overcome  the 
obstacles  incident  to  great  depth.  In  coal  mining,  longwall  has  a 
particular  advantage  in  thin  seams  over  the  pillar  system,  because  the 
robbing  of  pillars  in  such  seams  in  usually  unprofitable,  and  longwall 
reduces  to  a  minimum  the  expense  of  driving  and  maintaining  the  road- 
ways. Longwall  also  gains  in  desirability  with  increasing  depth  where 
the  pillars  of  the  rival  system  must  continually  widen  and  thus  proportion 
ately  be  dearer  to  recover. 

Longwall  is  adapted  to  beds  containing  considerable  waste,  for  the 
waste  can  all  be  stored  underground  and  if  suitable  for  pack  walls  will 
obviate  the  use  of  timber  cogs.  In  Europe,  with  plenty  of  waste  for 
gobs  and  packs,  seams  as  thick  as  10  ft.  have  been  worked  in  one  slice 
by  longwall.  Less  timber  is  consumed  in  the  longwall  than  in  the  pillar 
system  because  in  the  latter  the  props  can  seldom  be  used  again.  The 
subsidence  of  the  longwall  roof  is  gradual  so  that  it  does  not  inflict  such 
breaks  in  the  formation,  to  let  in  water  or  to  damage  surface  structure, 
as  ensue  from  pillar-robbing.  By  longwall,  a  mine  can  be  developed 
more  quickly  and  more  cheaply,  and  more  lump  coal  and  a  higher  per- 
centage of  the  seam  can  be  excavated  in  less  time  than  by  pillar  work. 
In  longwall,  the  ventilation  system  is  cheaper  to  construct  and  to  main- 
tain, for  the  mine's  resistance  is  less;  few  or  no  explosives  are  needed; 
and  there  is  less  dnger  from  falls  of  the  roof.  Longwall  requires  better 
trained  miners  than  pillar  work  but  a  miner's  output  is  greater.  After 
a  strike  of  nearly  six  months  recently  in  a  Western  coal  district, 
it  cost  the  pillar  mines  nine  times  as  much  to  clean  up  and  to  get  started 
again  as  it  did  similar  mines  where  the  longwall  system  prevailed.  This 
result  was  strictly  opposite  to  the  opinion  previously  held  by  many  on 
the  subject.  Longwall,  however,  is  unsuited  to  fluctuating  outputs,  for 
the  roof,  when  being  moved  at  all,  should  subside  uniformly  along  the 
face.  For  coal  mining,  longwall  is  gradually  superseding  pillar  work 
in  Europe  wherever  conditions  are  suitable.  America  is  bound  to  follow 
suit  as  soon  as  her  mines  become  deeper. 

(6)  COMPARISON  OF  THE  RETREATING  AND  ADVANCING  SYSTEMS 

The  retreating  system  of  mining  seams  is  rapidly  supplanting  its 
rival,  the  advancing  system,  in  European  mines,  but  in  America  the 
author  knows  of  no  case  of  its  use  in  longwall  and  of  comparatively  few 
cases  in  pillar  working.  On  inquiring  why  the  advancing  system  is  still 
preferred,  the  only  two  reasons  to  be  found  for  its  use  in  opening  a  mine 


230  MINING    WITHOUT    TIMBER 

are  that  it  requires  less  capital  and  less  time.  Everything  else  is  against 
the  advancing  system  which  has  a  smaller  percentage  of  mineral  recovery, 
a  worse  control  of  roof,  ventilation,  and  drainage,  a  greater  liability  to 
gas  and  dust  explosions,  a  higher  cost  of  maintenance  of  roadways,  a 
larger  timber  consumption,  and  a  smaller  output  from  an  equal  developed 
area  at  the  face. 

The  mistake  of  sacrificing  safety,  mineral  and  profit  per  acre  in  order 
to  get  an  output  quickly  at  minimum  cost  may  be  unavoidable  for  small 
weak  enterprises,  but  no  valid  excuses  can  be  made  by  strong  companies 
which  continue  to  persue  such  a  penny-wise,  pound-foolish  policy,  The 
driving  of  entries  to  the  boundary  to  inaugurate  the  retreating  system 
for  either  longwall  or  pillar  working  completely  explores  the  traversed 
territory.  These  entries  expose  the  seam's  faults,  rolls  and  irregu- 
larities, and  thus  indicate  both  the  lowest  points  of  the  floor  for  the  loca- 
tion of  sumps  and  pumps,  and  the  high  points  of  the  roof  where  may  be 
placed  churn-drill  holes  for  the  escape  of  gas  or  safety  shafts  with  ladders 
to  serve  as  natural  ventilators  when  the  fan  is  idle. 

With  the  retreating  system  not  only  are  there  no  old  gobbed  areas 
within  the  active  workings  to  generate  foul  gases  and  fires,  but  before 
stoping  begins  and  fills  up  the  mine  with  men,  the  seam  has  been  per- 
forated everwhere  by  the  entries,  and  most  of  its  water-channels  and 
pockets  or  feeders  of  gas  have  been  discovered  and  placed  under  control. 
In  retreating,  when  stoping  begins,  drainage,  ventilation  and  tramming 
are  covering  the  whole  area  of  the^  property,  and  are  at  a  maximum; 
and  all,  especially  the  two  latter,  tend  to  grow  less  as  the  working  area  is 
contracted,  while  the  advancing  system  implies  a  continual  extension  of 
the  area  covered  by  each.  The  maintenance  of  entries  is  a  serious  ex- 
pense in  an  advancing  system,  as  they  must  not  only  be  constantly  re- 
brushed,  but  are  liable  to  develop  irregularities  from  squeeze  which 
make  uniform  tramming  grades  difficult  to  maintain.  The  final  capital 
cost  of  entries  is  the  same  in  either  system  but  the  cost  of  maintenance 
for  retreating  is  only  a  fraction  of  that  for  advancing.  This  gain  alone 
will  often  more  than  offset  the  earlier  outgo  of  capital  requisite  for  the 
former  system. 

With  the  advancing  system,  coal  is  apt  to  be  lost  even  by  longwall, 
while  the  history  of  even  recent  pillar  working  in  America  indicates  that 
an  average  of  hardly  70  per  cent,  of  the  seam  is  recovered.  To  obviate 
the  only  two  drawbacks  to  retreating,  the  need  of  much  capital  and  time, 
the  two  systems  can,  in  pillar  working,  be  easily  combined  temporarily, 
by  opening  off  enough  rooms  from  the  advancing  entries  to  maintain  a 
modest  output  until  the  boundary  is  reached,  where  pillar-drawing  can 
then  be  started,  and  the  regular  retreating  system  inaugurated.  In 
longwall  working,  the  combination  of  advancing  and  retreating  is  less 
simple  because  of  the  complications  it  is  liable  to  cause  in  the  control  of 


PRINCIPLES    OF   MINING    SEAMS  231 

roof,  especially  with  the  continuous  face  method;  but  with  the  panel 
layout,  it  can  be  effected  in  those  cases  where  the  roof  is  flexible  enough 
to  permit  the  longwall  operation  of  isolated  panels  of  moderate  size. 

The  longwall  practice  cited  in  the  next  chapter  is  all  on  the  advancing 
system  owing  to  the  lack  of  retreating  examples  in  America,  but  the 
advance  layouts  described  can  readily  be  transformed  into  those  for 
retreat  by  merely  starting  the  initial  longwall  face  at  the  boundary  of 
the  property  instead  of  at  its  entrance. 

(c)  MINING  BY  ROOF-PRESSURE 

Blasting  must  ever  be  a  danger  in  a  colliery;  and  all  practical  sub- 
stitutes not  involving  the  creation  of  'flame  or  high-temperature  gases  are 
to  be  welcomed.  The  different  forms  of  wedges,  hydraulic  cartridges, 
lime  cartridges,  and  other  appliances  of  like  purpose  have  all  received 
full  attention,  but  little  has  been  written  on  Nature's  own  solution  of 
the  difficulty;  viz.,  roof  pressure,  and  its  systematic  and  scientific 
utilization. 

Any  bed  or  seam  is  subjected  to  a  certain  compressive  force  owing  to 
the  weight  of  the  superincumbent  strata :  if  a  portion  of  the  bed  is  removed 
and  no  artificial  means  of  supporting  the  excavation  attempted,  a  "  center 
of  relief"  is  established,  the  roof  and  floor  of  the  cavity  move  together, 
and  the  coal  (or  other  material)  round  and  about  the  cavity  is  cracked 
and  crushed  by  the  roof  weight,  and  eventually  some  of  it  forced  out  into 
the  open  space.  If  the  coal  surrounding  the  excavation  had  been  under- 
cut, it  would  have  fallen  under  the  action  of  the  roof  weight  sooner  and 
in  better  condition;  but,  had  the  undercut  been  too  deep,  the  coal  would 
have  fallen  en  masse,  and  have  necessitated  manual  labor  in  breaking  to 
a  size  .suitable  for  removal. 

The  cleavage  of  the  coal  must  be  studied  in  order  to  determine 
its  behavior  under  roof  pressure.  The  terms  bord  (or  face)  and  end 
(or  butt)  are  pretty  universally  employed  in  application  to  a  coal  face 
advancing  with  its  length  parallel  to  the  planes  of  main  cleavage  or  cleat, 
and  perpendicular  to  those  planes  respectively.  It  is  well  known  to 
every  collier  that  bordways  is  the  easiest  direction  of  advance;  but  coal 
so  hewn  is  most  likely  to  result  in  a  high  proportion  of  slack.  On  the 
other  hand,  the  coal  is  strongest  end-on;  is  hardest  to  hew,  but  is  most 
likely  to  result  in  a  large  percentage  of  lump  when  so  obtained. 

The  mode  of  fracture  of  the  roof  also  needs  attention.  The  forces 
which  induced  the  cleat  into  the  coal  had,  in  the  generality  of  cases,  a 
similar  effect  on  the  strata  above,  causing  an  incipient  cleavage  in  it 
coincident  in  direction  with  the  cleat  of  the  coal.  For  this  reason,  the 
maintenance  of  a  long  straight  face  absolutely  bord  is  almost  an  impos- 
sibility; at  such  a  face  the  roof  would  be  beyond  control,  would  break 


232 


MINING    WITHOUT    TIMBER 


off  "short"  against  the  face  (Fig.  108),  and  would  not  only  be  a  constant 
source  of  danger,  but  would  take  most  of  the  useful  weight  from  the  face. 
This  last  fact  was  recognized  very  early  by  coal  miners  and  wishing 
to  combine  the  easy  bord  direction  of  advance  with  a  better  control  of 
roof,  they  instituted  the  stepped  face  (Fig.  119).  Since  the  mean  line 
of  advance  in  the  case  shown  in  the  figure  runs  some  30  deg.  from  bord, 
it  follows  that  the  roof  will  break  parallel  to  this  line.  Stepped  long- 
wall  has  many  disadvantages;  first  among  which  must  be  placed  the 
fact  that  the  stepped  face  is  unsuited  to  machine  holing.  Secondly,  out- 
standing points  of  coal,  such  as  K,  Fig.  119,  receive  an  undue  roof  weight 
and  become  crushed.  While  at  the  other  extreme  we  have  points  such 


FIG.  108. — Effect  of  pressure  on  roof. 


as  m,  Fig.  119,  too  far  back  and  too  well  protected  for  the  roof  weight 
to  act  usefully  there,  where  also  the  coal  is  bound  on  two  sides  (along 
the  face  and  down  the  step),  and  correspondingly  hard  to  hew.  Again, 
since  the  packs  have  to  be  built  close  against  the  side  of  the  step  to 
support  the  coal  (the  space  between  the  two  is  seldom  more  than  2  ft., 
often  less)  the  ventilating  current  suffers  from  such  restrictions — an 
effect  which  is  further  augmented  by  a  frictional  loss  brought  in  by 
the  air  being  forced  to  travel  a  zig-zag  path.  Stepped  longwall  is 
giving  way  in  places  to  the  straight  "half-on"  face,  which  allows  of 
machine  holing  and  a  well-controlled  roof. 

Further  factors  which  must  receive  consideration  in  a  discussion  of 
the  effects  of  roof  pressure,  beyond  those  outlined  above,  are : 

1.  The  nature  of  the  seam. 

2.  The  nature  of  the  floor  and  roof. 

3.  The  rate  of  advance  of  the  face. 

4.  The  amount  of  dip  of  the  strata  and  the  direction  of  dip  as  com- 
pared with  the  direction  of  the  cleat. 

To  utilize  the  roof  weight  to  the  best  advantage,  the  coal  must  be 
undercut  to  a  certain  uniform  depth,  such  that  when  the  sprags  are 
withdrawn  the  coal  falls  with  a  vertical  fracture  from  the  back  of  the 
undercut,  without  any  extraneous  aid  by  blasting,  or  even  wedging. 

To  achieve  this,  the  undercut  must  generally  be  deeper  than  what  is 


PRINCIPLES    OF   MINING   SEAMS  233 

considered  advisable  by  hand;  hence,  we  must  depend  on  the  machine 
to  make  this  desideratum  an  actuality.  At  the  Altofts  Colliery,  Nor- 
manton,  Yorkshire,  they  have  succeeded  in  almost  dispensing  with  blast- 
ing by  holing  5  ft.  6  in.  under  in  a  flat  3  ft.  3  in.  seam,  1500  ft.  below  the 
surface.  With  a  hard  seam  a  much  deeper  undercut  than  51/2  ft.— 
perhaps  7  1/2  ft.  or  8  ft.  in  some  cases — would  be  found  necessary  if 
blasting  were  to  be  abolished;  such  a  depth  would  if  it  became  anything 
like  general,  cause  the  abandonment  of  disk  machines  in  favor  of  those 
of  either  the  bar  or  "puncher"  type. 

A  most  important  advantage  of  the  coal-cutting  machine  lies  in  the 
straightness  and  length  of  face  necessary  for  successful  application:  .the 
straightness  of  the  face  enables  the  timbering  to  be  absolutely  systematic ; 
and  this  factor  together  with  the  great  length  of  the  face  allows  of  the 
roof  pressure  to  be  controlled  and  utilized  with  precision.  Where  faults 
are  absent,  the  longer  the  longwall  face,  the  more  effectively  may  the 
roof  pressure  be  employed  as  the  means  of  breaking  down  the  coal. 

The  weight  of  the  roof  is  not  the  only  force  in  action  on  the  coal; 
before  the  coal  is  worked  the  pressure  of  the  floor  is  exactly  equal  and 
opposite  to  that  of  the  roof  on  the  seam;  when  an  excavation  is  made  in 
the  seam,  the  floor,  expanding  on  being  relieved  of  much  of  its  com- 
pression, exerts  an  upward  force  which  at  first  is  of  the  same  intensity 
as  the  roof  pressure  but  which  becomes  dissipated  sooner  than  the 
latter;  nevertheless  the  floor  pressure  is  often  of  service  to  the  miner 
and  is  taken  advantage  of  at  many  collieries  where,  owing  to  there  being 
a  suitable  band  of  dirt  at  or  near  the  top  of  the  seam,  overcutting  is 
resorted  to  in  lieu  of  undercutting.  Coal  so  obtained  often  is  in  better 
condition  than  coal  obtained  by  undercutting. 

A  treacherous  roof,  which  breaks  and  falls  immediately  the  weight 
comes  on  it,  rendering  timber  of  little  avail,  is  an  undoubted  evil.  Much 
may  be  done  in  the  way  of  palliation,  however,  by  quickening  the  rate  of 
advance  of  the  face,  and  proportionally  shortening  it  to  maintain  a 
uniform  output  (the  same  is  also  advisable  in  the  case  of  a  soft  seam) .  It 
has  often  been  found  effective  in  keeping  up  a  bad  roof  to  leave  a  thin 
strip  of  coal  against  the  roof.  The  device  seems  to  act  something  like  a 
plaster  on  a  wound:  it  has  an  effect  out  of  all  proportion  to  the  slight 
increment  of  strength  it  supplies :  its  action  is  to  prevent  the  slacking  and 
slipping  of  the  roof;  to  maintain  it  in  its  entirety. 

Just  previous  to  installing  machine  cutting  into  a  colliery,  experi- 
ments must  be  made  to  ascertain  the  depth  of  holing,  such  that  when  the 
sprags  are  withdrawn  the  roof  pressure,  aided  by  the  weight  of  the  coal 
undercut,  is  sufficient  to  break  off  the  coal  at  the  back  of  the  holing.  The 
result  arrived  at  will  be  somewhat  (6  in.  or  a  foot)  short  of  the  correct 
figure,  inasmuch  as  the  coal  face,  when  undercut  by  machine,  will 
advance  two  or  three  times  as  speedily  as  when  the  holing  is  done  by 


234  MINING    WITHOUT    TIMBER 

hand,  and  hence  the  roof,  not  having  the  time  to  weigh  so  heavily  will 
require  a  larger  surface  on  which  to  act. 

The  coal  when  the  sprags  or  wedges  are  removed  must  fall  not  in  a 
solid  block  (Fig.  109),  but  well  cleaved  (Fig.  110)  and  ready  for  immediate 


FIG.  109. — Undercut  coal,  fallen  en  masse. 


filling.  Should  the  coal  fall  as  exemplified  by  Fig.  109,  the  defect  can 
generally  be  remedied  by  turning  the  face  more  toward  bord,  or  by 
lessening  the  rate  of  advance  (the  former  method  in  preference),  and 
experiment  in  that  direction  should  be  made  at  once. 


FIG.  110. — Undercut  coal,  fallen  in  blocks. 


Judging  from  present-day  experience,  little  need  be  feared  on  the 
count  of  coal  cutting  when  the  depths  of  our  mines  become  excessive; 
indeed,  under  the  heavy  roof  pressures  then  in  action  the  use  of  explo- 
sives at  the  coal  face  is  likely  to  be  abolished,  and  the  depth  of  holing 


PRINCIPLES    OF    MINING    SEAMS 


235 


necessary  will,  if  anything,  be  less  than  that  at  present  in  vogue.  In 
very  deep  mines,  however,  it  has  been  found  that,  although  the  character 
of  the  seam  remains  the  same,  the  percentage  of  slack  increases  with  the 
depth.  Before  the  Royal  Commission  on  Coal  Supplies,  Mr.  Martin 
opined  that  an  increase  of  depth  from  1200  to  2400  ft.  would  result  in 
an  increase  in  the  percentage  of  slack  from  the  same  seam  of  5  per  cent.; 
and  at  Pendleton  Colliery,  while  the  coal  was  worked  at  depths  of  less 
than  2  500  ft.,  the  percentage  of  slack  was  21.5,  but  when  the  workings 
had  reached  the  depth  of  between  3000  and  3500  ft.,  the  proportion  had 
increased  to  39  per  cent.  This  is  merely  another  way  of  stating  that  the 


FIG.   111. — Roof  pressure  when  mining  to  rise. 

roof  pressure  has  been  too  severe  for  the  coal,  and  to  mitigate  such  an 
effect  the  coal  should  be  got,  as  far  as  possible,  by  machine,  working  end- 
on,  and  with  a  rapid  rate  of  advance. 

The  effects  of  dip  on  the  action  of  the  roof  pressure  is  important. 
In  a  working  proceeding  full  rise,  experience  tells  us  that,  other  things 
being  equal: 

1.  Hewing  is  easier. 

2.  Work  is  more  dangerous  (from  falls  of  roof  and  face) . 

3.  More  slack  is  produced  than  in  a  similar  working  in  a  flat  seam. 
Hewing  is  easier  for  the  reason  that  both  the  roof  pressure  and  the 

weight  of  the  coal  have  more  total  useful  effect  in  a  rise  working  than  in 
any  other.  In  Fig.  Ill,  by  means  of  the  undercut  A  B  a  wedge-shaped 
block  of  coal  A  B  C  D  is  undermined,  if  sprags  or  wedges  be  placed 
under  the  mouth  of  the  undercut,  the  triangular  block  A  D  E  is  still 
unsupported,  giving  us  at  once  the  reason  for  the  liability  to  falls  of  face 
in  such  a  working,  and  also  demonstrating  the  need  for  the  cocker  sprag 
(shown)  or  equivalent  means  of  supporting  the  face.  The  action  of  the 
roof  is  two  fold.  There  is  a  pressure  P,  acting  normally  to  the  plane 


236 


MINING    WITHOUT    TIMBER 


of  the  seam;  there  is  a  thrust  T,  acting  in  the  direction  of  dip,  tending 
to  make  the  roof  slide  over  the  face  toward  the  empty  space  behind  it. 

The  force  T  is  evidenced  in  the  fact  that  a  fracture  in  the  roof  of  a 
rise  working  "  gapes,"  owing  to  the  lower  side  having  moved  slightly  down 
under  the  influence  of  T.  Thus  it  is  that  falls  of  roof  are  more  prevalent 
in  rise  workings  than  in  any  other;  the  side  thrust  T,  not  only  quickly 
breaking  up  the  roof,  but  also  widening  the  joints  the  better  to  allow 
severed  slabs  to  fall. 

It  is  largely  to  this  side  thrust  that  we  owe  the  production  of  slack 
which  is  one  of  the  disadvantages  of  rise  working;  grinding  is  introduced, 
a  far  more  effective  slack  producer  than  mere  normal  pressure. 

The  resultant  action  of  the  forces  P  and  T  on  the  coal  may  be  best 
represented  by  the  single  force  R.  The  direction  of  R  cannot  be  accu- 
rately assigned,  but  it  lies  somewhere  between  the  normal  (P)  and  the 
perpendicular  (shown  dotted),  and  its  position  is  probably  somewhere 
as  shown;  its  magnitude,  by  the  parallelogram  of  forces  is  simply\/P2  +  T2. 
Considering  the  coal  acted  on  by  R  aided  by  the  weight  of  the  coal  itself, 
no  further  demonstration  is  needed  of  the  reason  of  the  ease  experienced 
in  working  coal  to  the  rise. 


FIG.   112. — Roof  pressure  when  mining  to  dip. 

Rise  working  would  be  rendered  safer,  and  less  slack  would  be  produced 
if  a  rapid  rate  of  advance  were  maintained,  and  to  compensate  for  the 
lessening  of  roof  pressure  which  would  result,  a  deeper  undercut  would 
be  necessitated.  Carefully  built  pack  walls  are  also  highly  advisable  in 
a  seam  liable  to  produce  slack,  and  thus  especially  in  rise  workings. 

In  the  dip  working  (Fig.  112),  the  action  of  T,  the  side  thrust,  is  much 
less  important;  the  tendency  is  there,  but  the  action,  so  far  as  the  grinding 
of  the  coal  is  concerned,  is  nil.  Also,  any  line  of  break  appearing  in  the 
roof  is  closed,  instead  of  opened,  by  the  slight  lateral  movement  of  the 
roof  over  the  gob  or  goaf:  hence,  working  to  the  dip  of  the  seam  is, 
generally  speaking,  the  safest  of  all  directions  of  working  the  coal.  As 
before  the  resultant  roof  pressure  acts  slightly  down-hill  (shown  at  R), 
causing  the  coal  in  this  case  to  be  difficult  to  hew.  * 

The  coincidence  or  noncoincidence  of  the  direction  of  cleavage  and 
the  direction  of  dip  is  a  factor  of  importance,  influencing  the  behavior  of 


PRINCIPLES  OF  MINING  SEAMS  237 

the  coal  under  the  roof  pressures.  If  the  directions  bord  and  dip  coincide 
a  rise  working  is  doubly  easy,  but  the  face  will  need  stepping  or  the  roof 
will  be  beyond  control;  also,  under  these  conditions,  dip  working  will  be 
facilitated,  and  generally  the  face  in  such  a  working  may  be  maintained 
straight,  owing  to  the  side  thrust  closing  the  jointings  in  the  roof.  On 
the  other  hand,  should  the  directions  end  and  dip  coincide,  the  easiest 
mode  of  advance  will  not  be  full  rise  but  in  some  direction  between  that 
line  and  the  strike  of  the  seam,  while  the  difficulty  of  working  directly 
to  the  dip  will  be  intensified.  Intermediate  between  these  extremes 
there  is  an  infinite  number  of  angles  at  which  the  directions  of  dip  and 
cleat  may  lie,  every  case  needing  special  consideration  and  experiment. 


CHAPTER  XIX 
ADVANCING  LONG  WALL  SYSTEMS  FOR  SEAMS 

EXAMPLE  49.— SPRING  VALLEY  BITUMINOUS  COLLIERIES,  BUREAU 

COUNTY,  ILLINOIS 

(See  also  Example  5.) 

Thin  Flat  Seam  at  500-/£.  Depth.  Advancing  with  Continuous  Face  by 
Scotch  System.  Loading  into  Cars, — The  Illinois  coal  reports  show  that 
over  5,000,000  tons  are  produced  in  the  longwall  field  or  about  12  per  cent, 
of  the  States'  total  output.  Most  of  the  longwall  mining  is  done  in  the 
prairie-like  counties  of  Bureau,  Grundy,  and  La  Salle.  Here  the  coal  seams 
are  remarkable  not  only  for  their  variety  and  quality,  but  in  their  free- 
dom from  horse-back,  faults,  and  other  irregularities,  which  are  encount- 
ered elsewhere.  The  mines  are  developing  a  3  1/2-ft.  seam,  called  com- 
mercially "third-vein  coal,"  which  is  350  to  500  ft.  beneath  the  surface. 
Overlying  the  seam  is  a  flexible  shale  4  to  9  in.  thick  and  underlying  it  is 
6  to  24  in.  of  fireclay. 

The  Scotch  system  extracts  all  the  coal  in  the  first  operation,  com- 
mencing at  the  periphery  of  the  shaft  pillar  and  mining  out  the  whole 
seam  toward  the  property  limits.  In  Fig.  113,  the  ideal  plan  of  the  lay- 
out of  a  Spring  Valley  mine  350  ft.  deep,  h  is  the  hoisting,  a  the  air-shaft, 
and  km  the  shaft  pillar,  600  ft.  square,  over  which  are  located  the  shaft 
house,  shops,  and  other  necessary  surface  structures.  To  open  out  the 
seam  for  longwalling,  the  pillar  is  first  cross-cut  by  the  two  headings  gn 
and  pq,  the  former  being  double-tracked  to  permit  the  handling  of  cars  to 
and  from  shaft  h.  Roads  are  uniformly  9  ft.  wide,  excepting  at  the  part- 
ings where  they  are  14  ft.  wide  to  hold  two  tracks.  These  peripheral 
headings  are  driven  from  the  points  g,  n}  p,  and  q,  and  by  widening  them 
inbye,  the  first  longwall  face  is  begun.  At  first  the  face  is  not  continuous, 
but  in  arcs  like  those  dotted  around  points  g,  n,  p  and  q,  but  as  the  arcs 
move  inbye  they  finally  meet  and  form  one  continuous  circle.  Soon  the 
face  has  advanced  sufficiently  to  allow  the  first  break  to  be  made  in  the 
roof  around  the  shaft  pillar,  so  that  thereafter  advantage  can  be  taken  of 
the  pressure  from  the  descending  roof  to  mine  the  coal  as  in  Fig.  114. 

Fig.  113,  shows  the  longwall  face  after  it  has  advanced  some  distance 
from  the  shaft-pillars  and  appears  as  the  circle  1-4-7-10.  The  face  is 
reached  through  the  main  roadways  1,  2,  3,  etc.,  by  which  it  is  divided 
into  approximately  equal  spaces.  The  essential  feature  of  the  Scotch 

238 


ADVANCING    LONGWALL    SYSTEMS   FOR   SEAMS 


239 


system  by  which  it  differs  from  the  other  continuous-face  longwall  system, 
"  the  Rectangular,"  is  the  turning  off  of  cross-roads  at  45-deg.,  angles  from 
the  main  directions  for  roads,  g-n  and  p-q.  The  road  e-3  was  formerly 
the  extension  of  gk,  but  its  present  position  makes  possible  a  gentler 
curve  for  haulage  into  gh  and  permits  the  use  of  the  40-deg.  frog  of  the 
other  45-deg.  turnouts.  The  layout  of  roads  is  based  on  a  room  about 
42  ft.  wide  at  the  face  (as  at  /)  which  corresponds  to  60  ft.  along  the  road 


Overcast      =  X 
7  Door  or  Curtain  =^~» 

FIG.  113. — Plan  of  Scotch  Longwall  system,  Illinois- 

c-4.  The  angle-roads  are  turned  off  at  225-ft.  intervals  from  a  cross-road 
as  e-3.  Experience  has  shown  that  a  room  can  not  exceed  a  length  of  225 
ft.  with  out  having  its  track  rebrushed  before  completion.  Hence,  the 
rooms  of  an  angle-road  like  cd  are  abandoned  as  soon  as  they  are  cut  off 
by  the  next  road  r-4.  Crossroads  e-3  and  y-5  are  1178  ft.  apart,  so  that 
angle-roads  c-d  and  d-g  can  each  be  840  ft.  long  to  contain  exactly  14 
rooms. 

The  direction  of  the  air  current  is  shown  by  arrows.  The  shaft  a  is  the 
downcast,  the  shaft  h  is  the  upcast;  roads  12,  2,  6,  and  8  are  intake,  and  the 
other  roads  are  return  airways  along  their  inbye  portions.  In  the  intake 


240 


MINING    WITHOUT    TIMBER 


roads,  double  doors  are  placed  at  t,  t',  y,  and  y'  while  single  doors  are  placed 
at  points  with  strong  draft  like  z,  z' ,  etc.,  to  hold  the  air  along  the  working 
face.  Fire-proof  burlap  curtains  are  used  in  rooms  where  the  draft  is  not 
strong  enough  to  force  them  up.  The  ventilation  is  excellent  and  easily 
regulated.  Trap  boys  are  stationed  at  the  double  doors  and  also  at  cross- 
roads where  collisions  are  liable  to  occur  from  trains  approaching  in  oppo- 
site directions. 


FIG.   114. — Roof  sinking  behind  Longwall  face. 

For  haulage  the  grade  is  nearly  level;  big  mules  are  used  on  the  main 
roads  and  small  ones  for  gathering.  At  No.  4  mine,  7  gathering  mules 
with  two  cars  apiece  haul  250  cars  daily  from  the  face  to  the  second  parting, 
thence  4  mules  haul  trains  of  30  cars  to  the  first  parting  whence  they  are 
hauled  in  similar  4-mule  trains  to  the  shaft  bottom.  The  total  output  of 
1000  tons  is  hauled  in  the  day  shift  by  about- 40  mules,  the  face  being 
about  a  mile  distant  from  the  shaft.  The  cars  are  of  wood,  weigh  1100 
lb.,  hold  2700  lb.,  of  coal  and  run  on  a  track  of  42-in.  gauge,  laid  with  16- 
Ib.  rails  except  at  the  bottom  of  the  shaft. 


FIG.  115. — Plan  of  packs  and  tracks  at  Longwall  face. 

Fig.  115  is  a  plan  at  the  face  showing  two  room  "  gateways,"  or 
room  roads  with  tracks  turned  off  at  45  deg.  from  the  "  Timbered  Branch 
Road."  Halfway  between  the  gateways  is  the  "  mark"  a  which  separates 
a  room  into  two  21-ft.  halves,  each  assigned  to  one  miner  who  both  mines 
and  loads  his  coal  into  a  car  on  the  nearest  track  at  c.  The  undercut  is 
made  by  hand  pick,  from  a  crouching  position,  in  the  floor;  when  the  lat- 
ter is  of  sandstone,  the  coal  itself  must  be  grooved.  If  the  roof  is  working 


ADVANCING   LONGWALL   SYSTEMS   FOR   SEAMS 


241 


properly,  which  is  ascertained  by  sounding  the  face  with  a  hammer,  the 
undercut  need  only  be  20  in.  deep.  Otherwise  a  depth  of  5  ft.  is  some- 
times necessary.  The  clay  cuttings  are  thrown  back  into  the  gob. 

The  packwalls  along  the  gateways  are  6  ft.  wide  and  do  not  approach 
nearer  than  2  ft.  to  the  face  in  order  not  to  obstruct  the  air-current. 
They  are  built  of  slate  brushed  from  the  roof  (see  Fig.  109),  by  a  hand 
pick  to  a  height  of  about  61/2  ft.  above  the  tracks.  The  roof  of  the 
haulage  roads  sinks  as  the  face  advances  and  must  be  continually  re- 
brushed  for  slate,  which  is  partly  stored  in  an  abandoned  room  and  partly 
hoisted  in  cars  to  the  extent  of  10  per  cent,  of  the  coal-car  hoist.  The 
rebrushing  of  the  roads,  the  repair  of  the  track  and  the  retimbering,  is 
done  by  company  men  on  day's  pay,  but  initially  the  miners  brush  the 
gateways,  lay  the  tracks,  build  up  the  packwalls,  and  set  the  props  in  their 
own  rooms  as  part  of  their  contract  price  per  ton  of  coal  loaded. 

The  undercut  is  held  up  by  sprags  (see  Fig.  116)  until  the  whole  42-ft. 
face  of  a  room  has  been  completed.  When  the  sprags  are  knocked  out 


45  Turn  from  Branch 
to  Boom  Koad 

FIG.  116. — Section  of  Longwall  face,  gateway  and  branch  road. 

and  the  undercut  coal  does  not  fall,  it  is  wedged  down  in  two  layers  by 
steel  wedges  starting  from  a  shear  made  at  the  breast  center  or  "mark." 
To  reach  the  car  from  the  mark,  the  coal  must  be  reshovelled  twice  in  the 
narrow  alley  between  props  and  face.  Lines  of  8-in.  by  4  1/2-ft.  props, 
about  3  ft.  apart,  are  set  along  the  face  (see  Fig.  114)  at  necessary  inter- 
vals and  some  of  these  are  not  recovered.  The  haulageways  are  timbered 
with  three-quarter  sets  which  in  certain  places  are  seen  to  be  cribbed  10  ft. 
high  above  the  caps  to  catch  up  roof-caves  that  broke  down  the  original 
timbering.  Steel  I  beams  are  used  for  caps  over  some  of  the  double 
track  partings  and  the  shaft  bottom  road  is  walled  with  masonry. 
Wooden  cogs  are  placed  at  acute  road  corners  as  c  (Fig.  113)  and  also  to 
replace  props  along  the  coal  face  where  the  roof  is  unusually  weak. 

Work  can  cease  on  part  of  the  coal  face  without  injury  to  the  balance 
if  care  be  taken  to  keep  the  whole  face  regular  and  convexly  curved,  for 
trouble  ensues  if  corners  or  concavities  are  allowed  to  develop.  A  fall 
or  creep  of  roof  at  the  face,  which  closes  up  the  space  between  gob  and 
coal,  may  occur  from  uneven  advances,  from  failure  to  build  up  the  pack 

16 


242  MINING    WITHOUT   TIMBER 

walls,  from  unusually  weak  spots  in  the  roof,  or  from  long  shutdowns. 
The  face  is  reopened  as  "yardage"  work  by  driving  the  gateway  into  the 
coal  and  turning  off  a  heading  along  the  face  with  a  2-ft.  rib  between  it 
and  the  old  gob. 

The  haulageways  are  simply  gateways  selected  because  they  occur  at 
the  proper  intervals  of  the  layout  of  Fig.  113.  Near  the  face,  rooms  are 
always  advancing  in  two  directions  which  intersect  each  other  at  a  45-deg. 
angle.  Thus  room  zf,  as  it  advances,  is  cutting  off  all  the  rooms  between 
r-4  and  zf  which  are  then  abandoned  and  their  occupants  assigned  new 
working  places.  It  is  always  arranged  to  give  an  experienced  miner  a 
21-ft.  face  to  work  for  himself,  but  occasionally  he  takes  a  green  hand  as 
helper  and  pupil.  There  is  little  gas,  though  two  fire-bosses  are  employed. 
The  mining  and  hoisting  is  all  done  on  day  shift.  The  output  of  coal  is 
about  21/2  tons  per  man  employed  above  and  below  ground. 

EXAMPLE  50. — MONTOUR  IRON  MINES,  DANVILLE,  PA. 

Thin,  Sloping,  Shallow  Beds.  Loading  into  Cars. — The  long-wall 
method  of  minin  g  when  introduced  into  these  mines  was  but  little 
used  in  this  country,  and  seldom  in  beds  as  thick  as  these,  with  breasts 
frequently  4  to  5  ft.  high.  Figs.  117  and  118  show  the  general  method. 

Levels  are  driven  90  ft.  apart,  and  the  face  of  each  gangway  should 
be  kept  in  advance  of  all  higher  gangways,  so  that  of  the  gangways  C, 
E,  and  H,  for  instance,  the  face  of  C  should  be  the  farthest  from  the  slope 
or  the  mouth  of  the  drift;  but  in  fact  this  is  seldom  done  here,  as  it- 
necessitates  the  outlay  of  large  capital  before  any  return  is  realized. 

The  gangways  are  driven  7  to  10  ft.  wide  and  5  1/2  to  7  ft.  high,  so 
a  man  or  mule  can  go  erect.  The  lowest  gangway  C,  Fig.  117,  is  known 
as  a  " fast-end"  gangway,  as  it  is  driven  entirely  in  the  solid,  while  E,  H, 
etc.,  are  " loose-end"  gangways,  with  but  one  side  in  the  solid  and  the 
other  side  formed  by  the  stowing.  The  face  of  the  gangway  C  should  be 
kept  far  enough  ahead  so  that  blasting  there  will  not  interfere  with  the 
workmen  in  the  breast  D,  and  the  same  consideration  should  determine 
the  distance  of  the  breast  F  from  the  face  of  E. 

To  facilitate  the  loading  of  the  ore  into  cars  from  chutes,  the  gang- 
ways are  so  driven  that  the  roof  of  the  bed  will  lie  at  the  top  of  the  upper 
rib,  Fig.  1 18,  and  to  secure  the  proper  gangway  height  the  bottom  rock 
must  be  taken  up.  In  doing  this  along  the  lower  gangway  C  a  drainage- 
ditch  is  left  upon  the  lower  side.  When  first  driven,  the  gangway  C  is 
timbered  upon  the  upper  side,  but,  as  settling  takes  place,  the  props  are 
usually  broken,  and  it  is  necessary  generally  to  renew  them  and  to  blow 
down  the  roof  of  the  gangway,  which  frequently  settles  sufficiently  to 
obstruct  the  haulage-way.  Often,  however,  the  stowing  becomes  so 
tightly  packed  in  settling  that  retimbering  is  unnecessary.  In  the  soft 


ADVANCING    LONG  WALL   SYSTEMS    FOR   SEAMS, 


243 


ore,  by  reason  of  the  creeping  of  the  bottom,  the  gangway-props  must 
sometimes  be  renewed  several  times.  Breasts  or  rooms  D  and  F  are 
turned  off  at  an  angle  of  35  to  45  deg.  with  the  direction  of  the  gangway, 
depending  on  the  dip  of  the  bed.  The  breasts  are  24  to  30  ft.  long,  and 
usually  there  are  five  breasts  in  a  tier  between  two  gangways.  The 
height  of  the  breast  varies,  with  the  nature  of  the  ore,  from  2  to  5  ft.  In 
the  hard  limestone-ore  are  three  streaks  of  ore  which  are  taken  out  if 
sufficiently  rich;  but  if  the  ore  is  lean  the  central  streak  alone  is  taken 
out,  with  just  enough  rock  to  allow  the  mine  to  work  his  breast.  The 
hard  limestone-ore  and  the  block-ore  have  to  be  blasted,  but  the  soft-ore 
is  scraped  out  in  the  form  of  mud. 


FIGS.  117  AND  118. 

FIG.  117. — Long  section  of  stope,  Montour  mine. 
FIG.   118. — Cross-section  of  stope,  Montour  mine. 

Each  breast  is  worked  by  a  miner  and  one  laborer;  or  two  miners  will 
combine  and  work  two  breasts;  and,  sometimes,  one  miner  and  two  or 
three  laborers  will  work  two  breasts.  The  miner  working  the  top  breast 
of  a  tier,  such  as  D4,  Fig.  117,  also  drives  the  gangway  E,  takes  up  the 
bottom  rock  to  give  sufficient  height  for  haulage,  piles  the  stowing  care- 
fully on  the  lower  side  of  the  gangway,  and  prepares  the  road-bed  for 
the  track-layers,  for  which  additional  work  he  is  paid  extra. 

A  ditch  is  not  left  along  the  loose-end  gangway,  as  the  waier  should 
drain  through  the  stowing  to  the  fast-end  gangway  C.  The  gob  is 
thrown  loosely  between  the  breast  and  gangway  below,  excepting  along 
the  chutes  G,  where  it  is  carefully  piled  to  support  the  roof.  A  chute  or 
gateway  G,  21/3  ft.  wide,  is  left  for  each  breast,  down  which  the  ore  is 
thrown  to  the  gangway  below;  it  is  sometimes  lined  with  boards,  but 
generally  a  carefully-built  dry  wall  of  gob  suffices.  The  ore  from  each 


244  MINING    WITHOUT   TIMBER 

breast  is  carried  to  the  chute  by  hand,  and  drawn  out  from  its  bottom 
into  cars  as  desired. 

There  is  usually  a  platform  at  the  chute  bottom  to  facilitate  the  load- 
ing; in  the  soft  ore  it  is  placed  directly  above  the  car,  but  it  is  nearer  the 
bottom  in  the  hard  ore,  and  sometimes  the  ore  is  simply  allowed  to  pile 
up  along  the  gangway.  When  necessary  to  prevent  the  air-current 
drawing  up  a  chute,  a  canvas  curtain  is  hung  loosely  over  its  mouth;  but 
ordinarily  only  the  last  five  inside  chutes  are  kept  open,  the  others  being 
boarded  up  and  filled  with  gob  when  the  gangway  E  has  advanced 
enough  to  receive  the  ore  from  the  next  higher  tier  of  breasts.  The 
gangways  E,  H,  etc.,  are  connected  with  the  fast-end  gangway  C  by 
"pitching  gangways"  K  driven  through  the  gob,  and  back  of  these  last 
the  gangways  E  and  H  are  abandoned,  so  that  the  fast-end  gangway  C 
is  the  only  one  kept  open  through  its  entire  length.  The  gangways  in 
the  soft-ore  are  timbered  and  lagged  on  sides  and  top,  but  in  the  hard 
limestone-ore  and  in  the  block-ore  it  is  generally  necessary  to  timber  the 
sides  only,  as  the  roof  is  of  good  slate  or  sandstone. 

The  props  are  placed  2  to  6  ft.  apart,  depending  upon  the  nature  of  the 
roof.  In  the  hard  fossil-ore  and  in  the  block-ore  the  breasts  are  not 
timbered,  excepting  when  necessary  to  protect  the  chutes,  as  the  gob 
fills  up  the  space  and  supports  the  top.  In  the  soft  fossil-ore  small  props, 
3  to  5-in.  dia.,  are  used  to  keep  up  the  top,  as  the  gob  does  not  fill  more 
than  one-third  of  the  vacant  space.  Heavy  timbers  are  usually  placed 
along  all  chutes.  The  timber  is  furnished  by  the  company,  but  the 
miners  set  it,  both  in  the  breasts  and  along  the  gangway,  and  as  it  is  cut 
on  company  property  it  is  cheap. 

All  general  work,  such  as  track-laying,  the  clearing  away  of  "falls," 
etc.,  is  done  by  miners,  detailed  for  each  separate  piece  of  work,  instead 
of  by  laborers,  and  for  such  work  the  miners  are  paid  per  diem.  All 
drilling  is  done  by  hand,  and  the  ventilation  is  secured  by  natural  draft 
through  chimneys.  In  the  drifts,  the  cars  are  either  pushed  by  hand  or 
hauled  by  mules  to  the  mouth,  while  in  the  slopes  the  mine-cars  are 
hoisted  to  the  top  by  second-motion  engines. 

The  mine-cars  are  2  ft.  deep,  4  ft.  long,  and  3  1/2  ft.  wide,  and  hold 
about  one  ton  of  ore.  The  gauge  is  30  in.  and  the  wheels  are  14-in. 
dia.  and  loose. 

Cost  of  Mining. — For  breast-work,  miners  are  paid  by  the  ton  and 
for  gangway-work  they  are  paid  tonnage  and  yardage. 

Tonnage  payments  depend  upon: 

(1)  The  nature  of  the  ore. 

(2)  The  height  of  the  breast. 
Yardage  payments  depend  upon: 

(1)  The  nature  of  the  ore. 

(2)  The  kind  of  gangway. 


ADVANCING   LONGWALL   SYSTEMS   FOR   SEAM_S  245 

Since  the  nature  of  the  ore  in  the  fossil-beds  and  the  height  of  breast 
vary  so  irregularly,  it  is  almost  impossible  to  give  exact  figures  so  that 
chiefly  ratios  will  be  given.  Upon  a  basis  of  $1  per  ton  for  mining 
block  ore,  the  following  are  the  prices  paid  during  the  past  15  years: 

Block  ore  per  ton $1.00 

Block  ore  per  yard,  fast-end  gangway 4.00 

Block  ore  per  yard,  loose-end  gangway 1 . 70 

Hard  fossil  ore  per  ton 0 . 95 

Hard  fossil  ore  per  yard,  fast-end  gangway 6.25 

Hard  fossil  ore  per  yard,  loose-end  gangway 2 . 40 

Unskilled  labor  per  day 0 . 73 

Soft  ore  costs  to  mine  one-third  to  one-half  the  above  hard  ore  prices. 

One  ton  per  day  for  each  man  working  in  a  breast  is  considered  an 
average  output  for  a  shift  of  10  hours.  The  miner  pays  his  laborer,  or 
laborers,  per  diem,  at  the  above  rate.  In  gangway-work  the  average 
rate  of  advance  was  15  ft.  per  month  for  loose-end  gangways  and  7  ft. 
per  month  for  fast-end  gangways.  Owing  to  the  many  conditions  affect- 
ing the  rate  of  advance  along  the  gangways,  it  was  necessary  to  employ 
a  system  of  " allowances"  in  payment  of  gangway-yardage  so  as  to 
equalize  as  nearly  as  possible  the  pay  of  gangway-miners. 

EXAMPLE    51. — BULL'S    HEAD    ANTHRACITE    COLLIERY, 
PROVIDENCE,  EASTERN  PA. 

(See  also  Examples  5  and  59.) 

Thin,  Sloping,  Shallow  Seam;  Panel  System;  Loading  into  Buggies 
on  Endless  Rope. — The  coal  property  is  about  1200  ft.  square,  and  a 
section  of  the  measures  just  above  and  below  the  seam  being  mined 
longwall  is  approximately  as  follows: 

20    ft.,  slate  and  soil; 
2  1/2  ft.,  fireclay; 

1  ft.,  bone; 

5  ft.,  coal  seam; 
40     ft.,  slate; 

2  ft.,  sandstone; 

6  in.,  slate; 

"30  -  in."  coal  seam; 
18     in.,  hard  slate; 
9     ft.,  soft  shale; 
18     in.,  hard  slate; 

3  1/2  ft.,  coal,  4-ft.  seam; 
90     ft.,  sandstone  and  slate; 

8    ft.,  coal,  Diamond  seam. 


246 


MINING    WITHOUT   TIMBER 


Below  the  Diamond  seam  occur  the  Rock  seam,  the  Fourteen-foot, 
and  the  Clark,  all  of  which  and  also  including  the  Four-foot  and  Dia- 
mond seams  had  been  worked  out  by  room  and  pillar  prior  to  begin- 
ing  to  mine  the  Thirty-inch  seam  by  longwall. 

Consequently,  the  rock  above  and  below  the  Thirty-inch  seam  was 
cracked  and  in  many  cases  out  of  place,  the  cracks  often  extending  to  the 
surface.  In  consequence  the  footing  for  props  was  most  insecure,  and 
although  the  cover  above  the  Thirty-inch  seam  was  only  about  75  ft., 
it  was  impossible  to  hold  it  by  timbering  and  it  would  have  been  prob- 
ably impossible  to  take  out  the  coal  by  room  and  pillar.  The  longwall 
method  of  working  as  developed  by  Supt.  Vipond  is  shown  in  Fig.  119. 
A  rock  slope  was  driven  up  from  the  Four-foot  seam  at  a  slight  pitch  so 


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FIG.  119. — Plan  of  Longwall  system,  Bull's  Head  colliery. 

that  empty  cars  can  be  hauled  up  the  pitch  by  mules.  From  the  head 
of  this  pitch  the  gangway  a  was  driven  31  ft.  wide  and  5  to  6  ft.  high, 
bottom  rock  being  taken  up  to  give  sufficient  height.  At  the  same 
time  the  parallel  airway  b  was  driven  and  ventilation  secured  by 
means  of  the  headings  shown  through  the  gangway  pillars.  The  airway 
b  connects  by  a  passageway  &'  with  a  ventilation  shaft  from  the  under- 
lying Four-foot  seam.  The  rock  obtained  in  driving  the  airway  is 
piled  in  walls  along  both  sides  of  the  airway.  The  rock  resulting  from 
taking  up  the  bottom  during  the  driving  of  the  gangway  is  built  into 
a  continuous  wall  c  along  the  lower  side  of  the  gangway  and  into  walls 
d  16  ft.  wide  along  the  upper  side  of  the  gangway.  Through  the 
upper  walls  d  are  passageways  e  which  are  9  ft.  wide  and  are  spaced 
125  ft.  between  centers.  These  passageways,  called  gateways,  have 
loose  walls  b  8  ft.  wide  on  each  side,  thus  making  the  total  width  of  the 
gateway  and  the  walled  space  25  ft.  The  gateways  are  driven  the  same 
height  as  the  gangway  for  a  short  distance  in  from  the  gangway  so  as 
to  provide  a  place  for  the  mine  car  to  stand  while  it  is  being  loaded  and 


ADVANCING   LONGWALL   SYSTEMS    FOR   SEAMS 


247 


out  of  the  way  of  traffic  along  the  gangway.  This  distance  depends 
upon  conditions,  but  is  usually  not  over  40  ft.  Above  this  point  the 
gateway  is  made  only  the  height  of  the  coal  and  the  overlying  slate, 
that  is,  about  36  in.  The  coal  is  overlaid  by  about  6  in.  of  slate  which  is 
always  taken  down,  and  it  is  this  which  furnishes  the  greater  part  of  the 
material  needed  for  building  the  pack  walls  along  the  greater  length  of 
the  gateways  and  along  the  face  as  will  be  described  later. 

The  method  of  opening  out  a  face  is  shown  at  A.  Strips  are  taken 
off  the  face  parallel  to  the  gangway  and  the  gangway  walls  dt  and  as 
soon  as  sufficient  width  is  secured  between  the  gangway  wall  d  and  the 
face  of  the  coal,  a  track  h  is  laid  as  near  to  the  face  as  possible  so  that  it 
will  not  interfere  with  the  work  of  the  miners:  This  track  has  a  gauge 


FIG.  120. — Single  winch,  Bull's  Head  colliery. 

of  2  ft.  3  in.,  is  laid  with  25-lb.  rails,  which  are  10  ft.  long.  The  rail 
sections  are  joined  by  two  fish-plates,  one  placed  on  each  side  of  the 
flange.  The  rails  are  held  together  by  iron  bridles  which  are  laid  directly 
on  the  bottom.  On  this  track  is  a  small  buggy  into  which  the  coal  is 
shoveled.  This  buggy  is  moved  by  an  endless  wire-rope  operated  from 
a  point  i  on  the  gateway  as  follows:  A  cast-iron  wheel  a,  Fig.  120,  18  in. 
in  diameter  and  having  a  groove  2  in.  deep,  is  held  in  a  wooden  frame  b. 
At  the  bottom  is  a  pointed  iron  c  fixed  on  the  frame.  This  rests  upon  the 
bottom  rock.  At  the  top  is  an  adjustable  pointed  round  iron  d  the  lower 
9  in.  of  which  is  threaded  so  that  by  means  of  the  nut  e  countersunk  as 
shown  in  the  frame,  when  a  wrench  is  applied  to  the  squared  portion 
above  the  thread  the  point  can  be  forced  up  against  the  roof  and  the 
frame  thus  held  securely  in  place.  A  3/8-in.  wire  rope  is  wound  two 
or  three  times  around  the  wheel  a  so  as  to  give  it  sufficient  grip  on  the 
wheel.  At  the  other  end  of  the  track  along  the  face  at  /  this  rope  passes 


248  MINING    WITHOUT   TIMBER 

through  an  ordinary  6-in.  iron  block  and  tackle  which  is  hooked  to  a 
chain  placed  around  the  prop.  One  end  of  the  rope  is  attached  to  the 
front  end  of  the  buggy  and  the  other  to  the  back  end  of  the  buggy.  By 
turning  the  handle  /  of  the  wheel,  the  buggy  can  be  moved  forward  and 
backward  along  the  face.  This  buggy  is  made  of  timber,  holds  about 
20  cu.  ft.,  and  one  side  is  20  in.  high  and  the  other  18  in.,  this  difference 
being  made  to  allow  room  for  loading  over  the  side.  The  wheels  are 
6  in.  in  diameter  placed  on  1  1/4-in.  axles.  In  the  bottom  there  is  an 
iron  plate  which  slides  in  and  out  sideways,  being  moved  by  a  handle. 
The  track  h  along  the  face  A  which  is  just  being  started  extends  out  over 
the  track  e  and  by  pulling  out  the  slide  in  the  bottom  of  the  buggy  the 
coal  is  dumped  from  the  buggy  into  the  mine  car  standing  on  the  track  e. 

The  coal  is  not  undercut,  and  in  general  doas  not  need  to  be  drilled 
or  blasted,  the  weight  of  the  cover  being  generally  sufficient  to  loosen 
the  coal  with  an  occasional  shot  when  the  roof  pressure  is  not  sufficient. 
Owing  to  the  broken  conditions  of  the  measures  due  to  the  mining  out  of 
the  underlying  seams  the  coal  in  many  places  is  loose  and  simply  needs  to 
be  picked  out.  Six  men  work  along  each  face,  three  miners  and  three 
laborers  who  load  the  coal  into  the  buggies.  The  face  is  worked  in  several 
sections  as  shown  at  B,  each  section  being  taken  out  for  a  certain  distance, 
about  12  yd.  depending  upon  the  ease  with  which  the  coal  can  be  broken 
down;  but  no  section  of  one  facets  allowed  to  get  far  ahead  of  any 
other  no  matter  how  easily  the  coal  can  be  mined.  By  the  time  section 
3  has  been  mined  out  the  coal  in  section  1  will  have  again  loosened  by 
the  weight  of  the  cover  and  can  then  be  taken  out  after  the  cogs  have 
been  built.  The  track  h  is  moved  near  the  face  after  each  section  is 
mined  and  a  row  of  cogs  is  kept  close  up  to  the  track.  Each  face  of 
coal  is  kept  about  40  ft.  in  advance  of  the  next  following  face.  The 
cogs  are  6  ft.  square  and  the  rows  are  8  ft.  apart  parallel  to  the  face  and 
12  ft.  apart  perpendicular  to  the  face.  These  cogs  are  built  from  the 
slate  overlying  the  coal,  and  as  it  comes  down  in  large  slabs  a  very  firm 
cog  is  formed.  The  space  between  the  large  rocks  is  filled  in  with  dirt 
and  a  perfectly  solid  cog  thus  formed. 

As  already  noted,  after  the  gateways  e  have  been  driven  in  full  height, 
a  distance  sufficient  to  allow  the  mine  car  to  be  placed  in  the  gateway  out 
of  the  way  of  the  haulage  on  the  main  gangway,  the  height  of  the  gate- 
way is  decreased  to  the  thickness  of  the  seam  and  overlying  slate,  that  is, 
about  36  in.  The  coal  is  moved  from  the  face  to  the  car  at  the  mouth  of 
the  gateway  by  means  of  a  buggy  similar  to  that  used  along  the  face  and 
already  described,  but  instead  of  the  winch  shown  in  Fig.  120,  it  is  moved 
by  means  of  a  double  winch  placed  at  the  point  h,  Fig.  1 19,  on  the  gateway 
where  the  height  is  decreased  to  the  height  of  the  seam.  This  winch, 
Fig.  121,  has  two  drums  a  and  a'  which  run  loosely  on  the  axle  6,  but  by 
means  of  the  clutches  c  and  c'  by  means  of  a  lever  not  shown,  either  drum 


ADVANCING    LONGWALL   SYSTEMS    FOR   SEAMS 


249 


may  be  made  to  turn  when  the  handle  d  is  moved.  If  the  load  is  too 
great  to  be  moved  by  turning  the  handle  d  the  winch  may  be  operated  on 
second  motion  by  means  of  a  pinion  e  attached  to  a  movable  axle  /. 
One  box  of  this  axle  at  g  is  loosely  bolted  to  the  framework  allowing  a 
little  play  of  the  axle,  while  in  box  h  is  an  elliptical  instead  of  a  circular 
hole  through  which  axle  /  passes  so  that  it  can  be  pushed  over  to  the 
dotted  position  /'  throwing  pinion  e  out  of  gear.  The  winch  is  set  on  a 
framework  of  timbers  one  end  of  which  rests  directly  on  the  bottom, 
while  the  other  end  is  let  into  a  groove  in  a  prop.  There  are  two  ropes 


FIG.  121. — Double  winch,  Bull's  Head  colliery. 


which  wind  upon  the  drums  a  and  a'.  One  of  these  is  attached  to  one 
end  of  the  buggy,  while  the  other  passes  to  the  upper  end  of  the  gateway, 
thence  through  a  small  6-in.  iron  block  and  tackle  k  fastened  to  a  prop 
by  a  chain  and  back  to  the  other  end  of  the  buggy,  or  by  means  of  guide 
pulleys  or  rollers  at  the  inby  end  of  the  gateway  the  rope  may  be  carried 
along  the  face  to  a  return  pulley  m  at  the  extreme  end  of  the  face  and 
the  gateway  buggy  taken  along  the  face  and  the  coal  brought  directly 
from  the  face  to  the  gangway. 

The  conditions  for  operating  the  longwall  system  of  mining  are  par- 
ticularly unfavorable,  for  the  bottom  is  badly  broken  and  a  stable  footing 
for  props  is  often  unobtainable.  At  the  face  of  one  of  the  gateways  at 
the  time  of  our  visit  the  bottom  had  dropped  away  entirely  from  beneath 
the  coal  leaving  the  coal  supported  only  by  contact  with  the  overlying 


250  MINING^WITHOUT   TIMBER 

slate.  The  top  is  also  badly  broken,  allowing  the  surface  water  to  enter 
the  mine  and  giving  a  roof  that  cannot  be  controlled.  This  roof  settles 
down  over  the  gangway  packs  about  2  ft.  so  that  while  the  gangway  is 
driven  about  6  ft.  high  it  is  only  about  4  ft.  high  after  the  workings  have 
settled.  Over  the  gateways  and  cogs  the  roof  settles  about  2  ft.  Thus 
far  an  output  of  1800  tons  per  foot-acre  has  been  obtained,  a  much 
better  yield  than  is  usually  obtained  in  anthracite  mining. 

EXAMPLE  52. — VINTON    BITUMINOUS    COLLIERY,    VINTONDALE,    PENN. 

Thin  Sloping  Seam,  800  ft.  Deep;  Panel  System;  Loading  into 
Pan  Conveyors. — Transporting  coal  from  the  working  face  to  main 
haulage  roads  by  means  of  mechanical  conveyors  is  a  comparatively 
recent  departure  from  ordinary  mining  methods.  This  system,  which 
was  first  introduced  in  England,  was  early  recognized  by  leading  opera- 
tors there  as  possessing  superior  advantages  over  the  usual  manner  of 
working,  especially  in  thin  coal  seams  where  the  roof  has  to  be  brushed 
to  allow  all  but  the  tiniest  cars  to  reach  the  longwall  face. 

The  coal  worked  at  Vintondale  is  bound  tight  to  roof  and  floor  and 
is  the  "  B "  or  Lower  Kittanning  seam,  42  in.  in  thickness,  which  lies 
on  a  pitch  of  8  per  cent.,  with  an  average  of  200  feet  of  cover.  The  coal 
is  of  a  soft  and  friable  nature,  free  from  slate  bands  and  bony  coal,  but 
interspersed  with  sulphur  pyrites,  which,  at  times,  cause  considerable 
annoyance  in  cutting  and  drilling.  The  bottom  is  a  mixture  of  coal  and 
fireclay,  while  the  roof  is  composed  of  from  8  to  12  ft.  of  black  slate, 
overlaid  with  sandstone.  The  slips  in  the  slate  are  well  marked,  and 
lie  at  an  angle  of  25  deg.  with  the  line  of  greatest  dip ;  the  longwall  face 
is  kept  normal  to  these  slips.  The  present  panel  modification  of  longwall 
mining  was  first  started  in  No.  3  mine  in  1900.  At  the  outset  cars  were 
run  around  the  working  face  and  loaded.  This  method  brought  only 
fair  results,  owing  to  the  necessity  of  using  small  cars,  steep  grades,  and 
difficulty  in  keeping  roadways  open. 

Arrangements  were  then  made  for  the  placing  of  a  conveyor  along  the 
face,  allowing  the  cars  to  be  run  under  the  head-end  to  be  loaded.  The 
first  conveyor,  which  was  made  entirely  of  wood,  was  a  cumbersome 
affair,  and  much  time  was  consumed  in  moving  it  laterally  along  the 
face  after  the  cut  had  been  loaded  out;  but,  after  a  year's  trial,  the 
results  obtained  were  so  gratifying  that  metal  conveyors  were  designed 
and  ordered,  and  preparations  were  made  to  employ  this  system  on  a 
much  larger  scale  (Fig.  122). 

The  metal  conveyor  consists  of  a  trough  or  pan,  made  of  sheet  steel 
1/8  in.  thick,  12  in.  wide  at  the  bottom,  18  in.  wide  at  the  top,  and  6  in. 
high,  set  on  strap-iron  standards  as  shown  in  detail  in  Fig.  123.  A  con- 
veyor is  made  up  in  sections  of  6-,  12-,  15-,  and  18-ft.  lengths,  connected 


ADVANCING    LONGWALL   SYSTEMS   FOR   SEAMS 


251 


together  by  means  of  1/2-in.  flatheaded  bolts,  countersunk.  The  front 
is  inclined  for  a  distance  of  45  ft.  to  allow  clearance  for  mine  cars  to  pass 
under  (see  Fig.  114).  The  rear  end  is  inclined  for  15  ft.  to  compensate 
for  the  size  of  sprocket  wheel.  A  return  runway  for  the  chain  is  afforded 
below  the  pans  by  angle  irons. 

A  cast-iron  driving  sprocket,  18  in.  in  diameter  and  13-in.  face,  is 


Section  C-D 
FIG.  122. — Plan  and  section  of  Vinton  conveyor  system  showing  head  of  main  conveyor. 

attached  to  the  front  end.  On  the  shaft  of  this  sprocket,  which  is 
extended  12  in.  beyond  one  of  the  bearings,  is  keyed  a  12-tooth,  16-in. 
diameter  sprocket,  which  connects  with  the  driving  mechanism.  The 
rear-end  section  (c)  consists  of  a  framework  made  up  of  two  I-beams,  6  ft. 
long  and  strongly  braced,  on  which  rest  the  take-up  boxes  for  keeping 
the  chain  in  adjustment,  and  the  rear  sprocket  wheel  over  which  the 


FIG.  123. — Cross  section  of  conveyor,  Vinton  colliery. 

chain  returns.  There  are  two  conveyor  chains,  held  apart,  the  width  of 
the  trough,  by  crossbolts  which  act  as  scrapers  to  replace  the  usual 
plates.  The  chains  are  of  steel  and  are  designed  for  quick  repairing. 

The  triple  conveyor  system  (Fig.  124),  was  finally  designed  and  in- 
stalled as  an  improvement  over  the  single  type.  In  laying  out  a  mine 
for  this  system,  the  main  entry  and  airway  are  driven  up  or  down  the 


252 


MINING    WITHOUT   TIMBER 


pitch,  and  cross-headings  are  driven  off  them  at  intervals  of  400  ft.  at 
such  an  angle  as  will  give  a  2-per-cent.  grade;  75-ft.  barrier  pillars  are  left 
on  each  side  of  the  main  entries.  The  cross-heading  is  driven  20  ft.  wide 
and  gobbed  on  the  lower  side.  The  air-course,  which  afterward  is  used 
as  the  panel  or  block  face,  is  driven  20  ft.  wide,  but  no  bottom  is  lifted; 
a  40-ft.  pillar  is  maintained.  Block  headings  are  run  perpendicular 
to  cross-headings  at  518-ft.  centers;  they  are  driven  18  ft.  wide,  with 
bottom  lifted  in  the  center  5  ft.  wide,  and  deep  enough  for  a  5-ft. 
clearance. 

When  the  block  is  ready  for  operation,  a  conveyor  350  ft.  long  is 
placed  in  the  block  heading,  and  along  the  face  of  the  air-course  on  each 


ajo 


FIG.  124. — Plan  triple  conveyor  system,  Vinton  colliery. 

side  is  placed  a  conveyor  250  ft.  long,  with  delivery  ends  directly  over 
the  main  conveyor,  one  being  5  ft.  in  advance  of  the  other.  Each  con- 
veyor is  driven  by  a  20-horsepower,  250-volt,  series-wound  motor,  en- 
cased in  a  sheet-iron  frame  mounted  on  steel  shoes,  so  as  to  be  easily 
moved. 

Airways  are  maintained  on  the  blocks  by  driving  two  places  slightly 
in  advance  of  the  block  face,  6  and  4  ft.  wide,  respectively,  with  a  10-ft. 
pillar  between.  The  first  place  acts  as  a  stable  for  the  machine,  and  is 
driven  by  the  machine.  The  airway  is  pick-mined,  and  one  man  manages 
to  keep  these  places  going  on  the  rear  end  of  both  blocks.  By  this  ar- 
rangement no  cribbing  is  necessary. 

The  blocks  are  worked  to  within  25  ft.  of  the  cross-heading,  when 


ADVANCING    LONGWALL   SYSTEMS   FOR   SEAMS  253 

the  conveyors  are  removed  to  another  block.  The  remaining  pillar  is 
brought  back  along  with  the  heading  stumps. 

The  power  is  carried  to  the  top  of  the  block  heading  by  a  00  wire. 
Here  are  attached  two  insulated  twin  cables,  one  to  furnish  power  to  the 
machines,  the  other  for  the  drives  and  hoist. 

The  cables  are  carried  down  the  block  heading,  one  on  each  side  of 
the  main  conveyor,  being  attached  to  it  by  means  of  malleable-iron 
brackets.  At  the  junction  of  the  conveyors  connections  are  made  with 
the  drives,  also  with  a  cable  that  is  attached  to  each  of  the  face  conveyors. 

Stations  are  established  50  ft.  apart  on  the  face  conveyor  cables,  to 
which  connections  are  made  with  the  short  cable  attached  to  the  longwall 
machines  and  electric  drills.  Switches  are  placed  at  the  head-end  of  the 
main  conveyor,  by  which  the  power  is  controlled. 

The  method  of  handling  the  cars  to  the  conveyor  is  simple.  A  side 
track  is  laid  300  ft.  long,  of  which  the  block  heading  is  the  center.  Con- 
nection is  made  with  the  main  track  at  the  lower  end,  and  a  cross-over 
switch  is  placed  directly  under  the  conveyor.  At  the  upper  end  of  the 
siding  is  placed  an  electric  hoist.  A  trip  of  14  cars  is  shoved  into  the 
empty  track,  and  the  rope  is  attached  and  the  trip  pulled  up  to  the  con- 
veyor. Signal  wires  are  hung  between  the  conveyor  and  the  hoist,  and 
and  as  each  car  is  loaded  the  trip  is  pulled  forward.  When  loaded, 
trip  is  dropped  on  the  loaded  siding,  the  rope  disengaged  and  attached 
to  the  empties. 

The  crew  operating  a  double  block  consists  of  17  men,  i.  e.,  block 
boss,  machine  runner  and  helper,  driller,  shooter,  two  conveyor  men, 
hoist  boy,  five  loaders,  and  four  timbermen.  Two  longwall  machines  of 
Jeffrey  or  Sullivan  make  are  used,  one  for  each  side,  although  one  ma- 
chine can  keep  up  the  work  in  case  of  emergency.  The  machine  men 
finish  cutting  one  block,  in  five  hours,  and  then  put  the  machine  in 
position  to  start  back  on  the  cut  and  move  over  to  the  next  block  and 
begin,  cutting.  They  are  followed  by  the  shooter  and  loaders. 

When  a  block  is  cleaned  up,  the  timbermen  move  up  the  conveyor. 

This  consists  of  setting  a  line  of  props,  called  the  line  row,  about  8  ft. 
apart,  and  a  distance  from  the  conveyor  equal  to  the  depth  of  the  under- 
cut. As  these  are  placed  the  old  line  row,  which  is  now  against  the  con- 
veyor, is  withdrawn.  The  pulling  jacks  for  moving  the  conveyor  are 
distributed  along  the  block  40  ft.  apart  and  placed  in  position. 

The  shot  firer  keeps  closely  after  the  machine,  and  is  through  shooting 
shortly  after  the  undercut  is  finished.  The  driller  then  starts  from  the 
far  end  of  the  block  to  drill  holes  in  the  new  face.  It  usually  takes  him 
about  two  hours  to  drill  the  entire  width  of  the  block. 

Each  loader  is  supplied  with  a  pick  and  shovel  and  a  piece  of  sheet 
iron  9  in.  wide  and  6  ft.  long,  which  he  attaches  to  the  conveyor  to  act 
as  a  sideboard.  As  each  loader  cleans  up  his  place  he  moves  forward  to 


254  MINING    WITHOUT   TIMBER 

the  head  of  the  line.     This  continues  until  the  coal  is  loaded  out,  which 
usually  requires  about  six  and  a  half  hours. 

When  cleaned  up,  the  drive  is  reversed  and  the  timber  which  has 
arrived  on  the  last  trip  is  run  through  on  the  conveyor  to  such  points  on 
the  block  where  it  is  required.  When  this  is  accomplished  the  power 
is  shut  off,  and  the  conveyor  is  moved  up  to  the  line  row.  This  lateral 
move  of  the  convey  or -requires  very  little  time,  very  seldom  exceeding  five 
minutes.  A  break  row,  consisting  of  two  rows  of  props  set  2  ft.  apart,  is 
now  placed  along  the  lower  side  of  the  conveyor.  These  props  are  set 
on  a  cap  piece,  placed  on  a  small  pile  of  slack,*  and  wedged  at  the  top. 
Two  break  rows  are  all  that  is  necessary  to  protect  the  block.  In  the 
meantime,  a  portion  of  the  crew  are  engaged  in  pulling  out  the  extra 
break  row.  This  is  the  most  hazardous  work  on  the  block,  and  is  given 
personal  attention  by  the  block  boss.  Axes  are  used  in  this  operation, 
and  about  75  per  cent,  of  the  props  recovered  are  practicallyuninjured. 

While  part  of  the  crew  are  employed  timbering,  the  rest  make  the 
necessary  connections,  and  go  along  the  conveyor  with  a  pump  jack  and 
level  it  up.  They  also  build  a  crib  at  the  head  end,  which  is  placed  to 
prevent  the  roof  from  breaking  over  into  the  block  heading.  All  the  dead 
work  is  taken  care  of  by  the  four  timbermen,  thus  not  hindering  the 
steady  flow  of  coal,  which  averages  150  tons  daily,  from  a  5-ft.  undercut. 

For  the  purpose  of  keeping  the  machinery  in  as  good  shape  as  pos- 
sible, a  skilled  mechanic  is  attached  to  each  mine.  He  assumes  charge 
in  case  of  an  accident  and  makes  necessary  repairs,  although  most  of  the 
breakdowns  are  easily  taken  care  of  by  the  block  boss  and  machine  man. 

In  the  starting  of  a  block  is  where  the  best  results  are  obtained,  as 
the  roof  requires  little  attention  until  about  100  ft.  have  been  extracted. 
It  then  begins  to  weigh  heavy  on  the  posts,  and  it  is  found  necessary  to 
carry  three  or  four  double  break-rows  with  cogs  in  anticipation  of  what 
is  called  the  "big  break."  This  usually  occurs  when  the  block  is  ad- 
vanced from  100  to  150  ft.,  although  in  several  instances  a  500-f  t.  face 
has  been  carried  up  200  ft.  before  the  overhanging  strata  broke.  After 
the  sand  rock  is  down,  only  two  break  rows  are  carried,  and  the  roof 
keeps  breaking  behind  the  last  row  as  the  face  is  extended. 

The  men  are  paid  day  wages  and,  as  they  become  accustomed  to  the 
work  and  machinery,  are  advanced  accordingly.  The  " block"  boss, 
as  an  incentive  to  secure  the  best  results,  is  paid  a  small  bonus  per  ton 
besides  his  regular  day  rate.  The  cost  averages  for  the  last  two  years 
show  that  block  coal  is  loaded  on  the  mine  cars  35  per  cent,  cheaper  than 
the  district  mining  rate  for  pick  work  with  loading  into  cars. 

The  above  conditions  prevailed  in  Nov.,  1907,  but  on  the  author's 
visit  in  Sept.,  1910,  he  found  the  longwall  system  superseded  by  the 
former  room  and  pillar  system  for  the  following  assigned  reasons.  1. 
The  frequent  breakage  of  the  conveyors  caused  a  very  irregular  output. 


ADVANCING    LONGWALL    SYSTEMS   FOR   SEAMS  255 

2.  The  miners  preferred  the  contract  payments  of  room  and  pillar  to  the 
time  wages  of  the  longwall  system.  3.  The  timber  consumption  was 
excessive,  because  many  of  the  props  could  not  be  recovered. 

None  of  these  disadvantages  are  irremediable,  and  the  cost  of  timber 
can  always  be  obviated  wherever  enough  slate  can  be  cheaply  got  from 
the  floor,  parting,  or  roof  to  build  pack  walls  to  replace  part  of  the  props 
and  cogs.  The  conveyor-longwall  system  has  proved  profitable  for  thin 
seams  in  Europe,  and  the  Vintondale  method  should  prove  commercially 
successful  in  other  American  fields  where  the  natural  conditions  are 
suitable. 

EXAMPLE  53. — DRUMMOND   BITUMINOUS  COLLIERY,  WESTVILLE,   N.  S. 

Thick  Sloping  Seam  at  2000-/L  Depth.  Loading  into  Cars  handled 
on  "Jigs. " — When  coal  workings  extend  beyond  a  vertical  depth 
of  1500  ft.,  it  generally  becomes  unprofitable,  if  not  impossible,  to 
work  by  one  of  the  "pillar"  methods,  for  the  enormous  weight  of  the, 
overlying  strata  will  not  only  break  and  crush  the  timber,  but  will  also 
either  crush  the  pillars  or  force  them  into  the  strata  immediately  above 
or  below  the  seam,  resulting  in  a  "creep"  and  the  closing  up  of  roads. 

The  size  of  the  pillars  must  increase  with  the  depth,  until  at  about 
the  depth  noted  above,  the  pillars  become  so  large  and  the  amount  of 
coal  that  can  be  safely  worked  so  small,  especially  if  it  is  of  a  friable 
nature,  that  the  operations  become  unprofitable,  and  another  method 
must  be  adopted  or  the  mine  closed  up. 

Such  was  the  situation  in  1896  at  this  mine  in  working  by  the  room 
and  pillar  method  a  17-ft.  seam  of  friable,  gaseous  coal,  with  a  very  weak 
roof  of  black  carboniferous  shale,  the  seam  dipping  from  18  to  27  deg. 

The  mine  had  been  developed  by  two  parallel  slopes  on  the  dip,  and 
from  these  double-entry  "lifts"  were  turned  off  every  400  ft.  These 
entries  or  levels  were  9  ft.  wide  by  7  ft.  high,  and  as  depth  was  gained  it 
was  found  difficult  to  support  their  roofs.  At  a  depth  of  1200  ft.,  the 
first  "chocks"  or  cogs  had  to  be  built  on  each  side  of  the  level.  In  the 
next  lift,  400  ft.  below,  it  became  necessary  to  change  the  working  system 
if  the  coal  were  to  be  mined  at  a  profit. .  The  change  was  made  without 
great  expense,  any  interruption  of  the  regular  output  or  any  considerable 
variation  in  the  ventilation,  etc. 

The  advancing  panel  system  of  longwall  was  adopted  and  it  has  proved 
quite  successful  considering  the  depth  reached,  which  is  7,870  ft.  on  the 
slopes  or  over  2000  ft.  vertically.  The  new  system  has  been  partic- 
ularly free  from  fatal  accidents  at  the  face,  those  occurring  happening 
in  the  roadways,  etc.  The  slopes  are  sunk  as  formerly,  diverging  slightly 
to  increase  the  pillar  of  solid  coal  between  them.  They  are  supported 
on  either  side  by  pillars  also  increasing  in  width  with  depth  so  that  they 
are  now  about  350  ft.  wide. 


256 


MINING    WITHOUT    TIMBER 


Either  one  or  both  of  these  slopes  is  used  as  the  intake  airway,  Fig. 
125,  while  return  airways  are  maintained,  one  on  each  side  along  the  slope 
pillars.  Two  levels  are  driven  as  formerly  which  form  a  lift  with  about 
400  ft.  of  solid  coal  between  pairs  of  levels.  The  upper  level  of  each  pair 
is  used  as  a  haulage  road,  and  the  lower  level  forms  the  intake  airway 
for  each  lift.  This  intake  carries  fresh  air  from  the  slope,  where  it  is 
split,  to  the  inner  workings  first;  from  there,  returning  and  ascending,  it 
passes  through  each  of  the  working  places  to  the  lift  above,  and  thence  to 
the  return  airway;  it  is  also  used  for  drainage,  and  generally  there  is  a 
dam  built  on  it  near  the  slope  which  catches  all  the  water  from  the  lift. 


Air 


FIG.  125. — Plan  of  layout,  Drummond  colliery. 

The  levels  are  driven  as  nearly  parallel  as  possible,  rising  about  1  ft. 
in  130  ft.  with  from  15  to  20  ft.  of  solid  coal  between  the  chocks.  This 
pillar  is  often  removed  and  the  space  filled  in  with  stone  from  the  roof, 
the  result  of  "brushing"  which  must  be  done  very  shortly  after  the  levels 
are  driven.  These  levels  are  driven  8  ft.  wide  and  8  ft.  high,  they  are 
first  made  about  18  ft.  wide  and  7  ft.  high.  This  leaves  a  "bench"  on 
the  bottom  which  is  only  cut  in  the  case  of  roadways.  On  this  bench 
chocks  are  built  quite  close  together  on  each  side,  and  about  8  ft.  apart 
across  the  road,  Figs.  126  and  127,  with  sided  timber  over  them  across 
the  road  about  3  ft.  apart,  and  slabs  over  the  timber  to  support  the  roof. 
The  chocks  are  built  of  blocks  of  wood  over  5  in.  in  thickness  and  5  ft.  in 
length,  making  them  5  ft.  square.  After  these  are  built  (similar  to  logs 
in  a  wharf)  the  bench  is  cut  along  the  chocks  and  the  bottom  lifted  to 
give  the  8-ft.  height. 

Off  these  levels,  "jigs"  are  driven  up  on  the  full  pitch  of  the  coal, 
Figs.  127  and  128,  not  more  than  400  ft.  apart;  they  are  chocked  as  well 
as  all  other  roadways.  An  airway  5  ft.  wide  is  carried  up  on  the  side  of 
this  jig  farthest  from  the  slope,  and  the  chocks  on  this  si  do  must  be  made 


ADVANCING    LONGWALL    SYSTEM    FOR    SEAMS 


257 


air-tight.  This  is  done  by  filling  them  with  stone  and  fine  coal,  etc. 
Owing  to  the  very  heavy  pressure  required  in  maintaining  ventilat'on 
at  this  depth,  canvas  doors  can  only  be  used  as  a  temporary  arrangement. 
A  wooden  door  is  placed  in  an  air-tight  frame  across  the  level  to  direct 
the  air  up  this  airway  between  the  coal  and  the  air-tight  chocks,  passing 


FIG.  126. — Cross  section  of  road,  Drummond  colliery. 

around  the  face  and  returning  down  the  jig  which  is  8x8  ft.,  Fig.  127. 
This  practice  has  been  proved  many  times  to  be  the  only  practical  way, 
as  the  air  will  not  pass  up  the  large  and  down  the  smaller  airway  in 
sufficient  quantity  to  keep  the  face  clear  of  gas.  This  method  is  con- 
tinued until  the  jig  is  driven  through  to  the  lower  level  of  the  lift  above, 


A   Air  Tight  Chocks 
S    Stopping 


FIG.  127. — Plan  of  gateway  and  face,  Druminond  colliery. 

when  the  door  is  removed  and  the  air  passes  up  the  jig  and  out  to  the 
airway,  and  the  airway  along  the  chocks  is  allowed  to  cave. 

Working  the  Rooms. — Beginning  at  the  lower  entry  of  the  lift  above 
on  one  of  these  j:gs,  rooms  are  broken  off  with  about  41  ft.  between  the 
centers.  The  "gateway"  or  road  in  the  rooms  is  much  the  same  as  the 
levels  already  described.  They  are  timbered  in  the  same  way,  Fig.  128, 

17 


258 


MINING    WITHOUT   TIMBER 


except  that  the  chocks  are  built  about  2  ft.  apart  on  each  side,  and  only 
about  6  ft.  apart  across  the  road.  When  this  gateway  is  driven  in  about 
25  ft.,  work  is  started  on  the  breast.  From  it  the  coal  is  all  taken  out 
to  a  thickness  of  7  ft.  and  up  to  the  room  or  level  above.  The  breast  is 
then  timbered  with  upright  timber  props  with  cap-pieces  between  them 
and  the  roof.  Sometimes  sided  timbers  are  placed  with  one  end  on  the 
high-side  chock  of  the  roadway  and  props  under  the  middle  and  upper 
end.  The  gateway  is  kept  15  ft.  to  20  ft.  ahead  of  the  breast,  the  roof 
of  which  is  allowed  to  fall  in  as  the  face  advances;  generally  when  about 
40  ft.  from  the  jigs  the  roof  falls,  often  causing  a  great  smashing  of 
timber  on  the  road  below,  the  bottom  rising  up  as  well.  The  face  of  the 


Fid.  128. — Plan  of  jig  road,  Drummond  colliery. 

gateway  is  kept  a  short  distance  ahead,  for  if  the  face  of  road  were  in 
line  with  the  face  of  the  breast,  it  would  be  very  apt  to  fall  solid  across  the 
face  of  the  road  as  well,  and  take  a  week  or  more  to  get  into  working 
shape  again.  Through  carelessness  of  the  miners  this  sometimes  happens. 
No  explosives  are  used,  for  if  these  places  are  properly  timbered  and  the 
weight  thrown  on  the  face  the  coal  is  easily  worked  with  hand  picks,  but 
wedges  are  required  in  lifting  the  bench  in  the  roadways. 

Quite  often  the  roof  falls  in  solid  to  the  face,  then  it  is  necessary  to 
drive  a  heading  up  in  the  solid  coal  at  the  face  and  start  the  breast  over 
again.  This  perhaps  is  the  greatest  difficulty  met  with  in  the  whole 
operation,  for  when  a  fall  like  this  takes  place  the  ventilation  is  cut  off 
and  generally  some  gas  accumulates  and  when  the  heading  is  started  up, 


ADVANCING    LONGWALL   SYSTEMS   FOR   SEAMS  259 

the  gas,  also  rising,  follows  the  miner  and  causes  trouble  before  it  can  be 
driven  20  ft.  or  35  ft.  to  the  place  above. 

Generally  three  of  these  rooms  are  worked  on  each  side  of  the  jig 
simultaneously,  the  upper  ones  leading  and  the  others  following  in  step- 
like  order  from  20  ft.  to  40  ft.  behind  the  preceding  one.  In  this  way 
the  upper  7  ft.  of  the  17-ft.  thickness  of  coal  in  the  seam  is  taken  out  in 
one  operation  and  in  future  years  the  balance  may  be  mined  similarly. 

Three  miners  and  a  laborer  work  in  each  room,  as  the  success  of  this 
method  requires  that  the  breasts  be  kept  steadily,  if  slowly,  advancing. 
With  depth  it  is  necessary  to  shorten  the  gateways  in  order  that  the  coal 
may  be  all  taken  out  before  they  become  entirely  closed  up,  for  on  every 
side  may  be  seen  examples  of  both  squeeze  and  creep.  The  roof  pres- 
sure is  so  great  that  thin  clay  partings  in  the  coal  squeeze  out  like  clay 
from  a  brick-machine,  while  the  lateral  pressure  on  the  coal  walls  re- 
duces the  space  of  the  openings  30  per  cent,  in  a  few  months.  The 
combined  pressures  so  break  ordinary  booming  in  about  a  month  that  it 
becomes  necessary  to  brush  and  retimber  again.  Here  places  may  be 
seen  so  closely  timbered  that  neither  rock  nor  coal  are  to  be  seen  for 
long  distances  except  at  the  working  faces.  Studies  are  constantly 
made  of  the  faults,  cleats  and  clay  partings  encountered,  (so  as  to  make 
the  pressure  mine  the  coal  with  the  least  labor),  of  the  proper  setting 
of  timbers,  of  the  breaking  away  of  the  ribs  and  how  to  avoid  it;  and  of 
how  to  distinguish  the  actual  sounds  of  danger  as  there  is  always  some 
cracking  of  timbers  heard  in  the  working  places.  So  expert  do  those 
extracting  the  coal  from  the  breast  become  that  they  work  on  to  the 
last  minute  before  a  fall  takes  place  amid  appalling  conditions. 

Haulage. — There  is  no  mechanical  haulage  on  the  levels,  the  work 
being  done  by  horses.  At  the  bottom  of  the  jigs  and  at  the  mouths  of 
the  rooms — which  are  opposite  each  other,  three  on  each  side  of  the  jig — 
large  metal  plates  are  laid  on  timbers  placed  horizontally  and  made 
solid  in  that  position.  When  laid  they  form  a  smooth  surface  from  6  to 
8  ft.  square.  On  these  the  cars  can  easily  be  turned  in  any  direction. 

The  road  on  the  jig  is  either  a  double  track  or  three  rails,  with  a 
passing  turnout  halfway  (see  Fig.  128).  Over  the  two  lower  plates  on 
the  jig,  lifting  rails  about  8  ft.  long  are  fitted  into  clips.  When  running 
coal  from  the  lower  rooms,  these  rails  are  removed  and  a  tail-rope  the 
necessary  length  is  attached  with  a  safety  hook.  A  drum,  controlled  by  a 
boy,  is  placed  at  the  top  and  the  weight  of  the  full  car  running  down  takes 
the  empty  car  up.  The  cars  run  on  their  own  wheels  from  the  surface 
to  the  face,  their  gauge  is  2  1/3  ft.,  and  their  capacity  is  1600  Ib.  of  coal. 

Output  and  Timber  Used. — No  timber  is  drawn,  the  great  difficulty 
being  to  get  enough  timber  in  to  keep  sufficient  room  for  the  proper 
ventilation  of  the  mine.  The  mine  produces  about  1200  tons  of  coal 
a  day  and  consumes  two  thousand  5-ft.  sticks,  but  of  a  small  diameter. 


CHAPTER  XX 
PILLAR  SYSTEMS  FOR  SEAMS 

EXAMPLE  54. — ADVANCING-SYSTEM  LAYOUTS  FOR    ROOM  AND  PILLAR, 
PILLAR  AND  STALL,  AND  PANEL  METHODS 

Room  and  Pillar  System. — The  pillar  system  is  also  known  as 
"room  and  pillar, "  " pillar  and  chamber/'  "bord  and  pillar/'  etc.  It 
is  applicable  to  all  classes  and  conditions  of  mining  where  the  roof  pressure 
is  not  such  as  to  destroy  pillars  of  reasonable  sizes,  subject,  however,  to 
such  modifications  as  serve  to  adapt  it  to  the  varying  conditions  of  weak 


FIG.  129. — Layout  for  Room  and  Pillars  system,  flat  seam 

or  strong  roof  or  floor;  tough,  friable,  or  gaseous  coal;  predominance  of 
face  or  end  cleats;  inclination  of  the  seam,  etc.  The  features  of  the  sys- 
tem are  openings  driven  square  from  or  at  an  angle  to  the  haulway. 
Such  opening  may  be  driven  wide  or  narrow,  and  may  be  a  roadway, 
incline,  or  chute,  as  best  adapted  to  the  existing  conditions. 

Room  and  Pillar  System  (proper). — This  layout  as  applied  to 
flat  seams,  or  where  the  inclination  does  not  exceed  3  deg.,  is  illustrated 
in  Fig.  129.  The  shaft  bottoms,  including  the  stable,  are  here  shown 

260 


PILLAR   SYSTEMS   FOR   SEAMS 


261 


crossing  the  shaft  pillar  at  an  angle  conforming  to  the  surface  tracks, 
thereby  giving  a  straight  dump  and  tipple  in  line  with  the  shaft.  The 
stables  are  located  close  to  the  shaft  bottom,  where  the  mules  can  be 
rescued  in  case  of  accident,  and  where  the  daily  feed  and  refuse  can  be 
conveniently  handled.  Free  access  is  had  to  the  stables  from  the  main 
haulage  roads  without  passing  through  a  door;  while  immediate  access  is 
had  close  to  the  shaft  through  a  curtain  or  canvas.  Good  ventilation  is 
secured  by  a  small  separate  split  of  fresh  air,  while  the  return  air  from 
the  stables  at  once  enters  the  return  from  the  mine  and  passes  up  the 
shaft  without  contaminating  the  mine  air.  Another  feature  of  the 
arrangement  shown  in  Fig.  129,  is  the  small  number  of  doors.  The  coal 
coming  from  any  room  upon  the  main  road  of  a  pair  of  entries  has  no 


FIG.  130. — Layout  for  Room  and  Pillar  system,  sloping  seam. 

doors  to  pass  through;  while  that  coming  from  the  back  entry  of  each 
pair  has  but  one  door  to  pass  through  on  its  way  to  the  shaft.  This  is  a 
great  saving  of  expense  and  trouble  and  may  often  avert  possible  disaster 
arising  from  the  derangement  of  the  ventilation  by  doors  being  left  open. 
The  chambers  or  rooms  are  here  turned  square  with  the  entry  narrow 
for  a  distance  of  4  or  5  yards  and  then  widened  out  inbye,  the  road  in 
each  room  following  the  straight  rib.  The  waste  from  the  seam  is  stored 
in  the  room.  The  rooms  are  spaced,  under  normal  conditions  of  roof 
and  floor,  from  40  to  45  feet  apart,  center  to  center.  The  breast  is  usually 
8  yards  wide  and  is  driven  up  from  60  to  100  yards.  When  the  breast  is 
abandoned  the  miner  starts  to  draw  back  his  pillar  unless  for  special 
reasons  this  is  delayed  for  a  while. 


262 


MINING    WITHOUT    TIMBER 


In  Fig.  130  is  shown  the  application  of  the  room  and  pillar  system  to 
seams  pitching  from  3  to  5  deg.  It  differs  from  the  method  shown  in 
Fig.  129  by  turning  rooms  to  the  rise  only.  When  the  pitch  of  the  seam 
is  from  5  to  10  deg.,  the  car  may  still  be  taken  to  the  face  and  loaded  by 
driving  the  rooms  across  the  pitch,  or  at  an  angle  with  the  level  or  gang- 
way. This  reduces  the  grade  of  the  track  in  the  rooms.  When  the 
inclination  of  the  seam  is  still  greater,  buggies  are  sometimes  used,  the 
track  being  built  upon  the  refuse  of  the  seam  and  raised  at  its  lower  end 
where  a  tip  is  arranged  by  which  the  coal  is  dumped  from  the  buggy  into 
the  mine  car  ready  to  receive  it.  Where  the  coal  is  soft,  this  method 
cannot  be  used.  It  is  employed  on  pitches  not  exceeding  15  or  18  deg. 
in  thick  seams. 

Seams  pitching  more  than  15  deg.  are  usually  worked  by  chutes 
or  self-acting  inclines.  When  the  pitch  is  less  than  30  deg.  sheet  iron  is 


FIG.  131. — Room  and  Pillar  system  for  steep  seam. 

usually  laid  in  the  chute  as  a  floor,  to  enable  the  coal  to  slide  more  easily; 
but  on  inclinations  of  less  than  20  deg.  it  is  usually  necessary  to  push  the 
coal  down  the  chute  by  hand  or  by  mechanical  means,  as  it  does  not  slide 
readily.  On  pitches  steeper  than  30  deg.  sheet  iron  is  not  necessary,  as 
the  coal  will  slide  without.  Fig.  131  shows,  in  plan  A  and  section  B, 
the  arrangement  of  the  breasts,  chutes,  and  manways,  and  the  position 
of  the  gangway  and  air-course  at  the  roof  of  the  seam,  in  thick,  steep- 
pitching  anthracite  beds.  This  position  of  the  gangway  and  air-course 
secures  a  better  inclination  of  the  loading  chute  and  manways,  and  pre- 
sents less  danger  from  squeeze.  In  the  figure,  g  is  the  gangway  and  m 
the  manway  leading  to  the  dividing  at  the  floor  of  the  seam  into  two 
branches  s,  s,  which  lead  to  either  breast.  At  this  point,  or  slightly 
below  it,  a  small  cross-cut  d  is  driven  up  to  the  airway  c.  This  is  brat- 
ticed  off  and  used  only  in  case  of  need,  as  the  air  is  regularly  conducted 


PILLAR   SYSTEMS   FOR   SEAMS 


263 


up  one  breast  manway  and  down  the  other  side  to  the  highest  cross-cut 
and  thence  to  the  next  breast.  Brattices  with  small  doors  are  also 
placed  in  the  manways  to  keep  the  air  from  taking  a  short  circuit  through 
the  manways.  Small  manways  are  bratticed  off  the  side  of  each  loading 
chute  for  the  use  of  the  loaders. 

Self-acting  inclines  are  used,  sometimes,  upon  steep  pitches  in  pref- 
erence to  chutes.  In  this  case,  butt  headings  are  usually  driven  to  the 
full  rise  and  rooms  set  off  on  the  strike  from  these  rise  headings,  buggies 
being  used  in  the  rooms  to  convey  the  coal  from  the  face  to  the  incline. 
It  is  hardly  necessary  to  state  that  dip  inclines  are  rarely  ever  introduced 
as  a  permanent  feature,  it  being  better  to  sink  the  main  slope  far  enough 
to  permit  another  level  from  which  the  coal  can  be  worked  to  the  rise. 


FIQ.  132. — Single  Stail  system. 


FIG.  133. — Double  Stall  system. 


Stall  and  Pillar  System. — This  is  similar  to  the  system  just  described, 
except  in  the  relative  size  of  pillars  and  breasts.  It  is  adapted  to  weak 
roof  and  floor,  or  strong  roof  and  soft  bottom,  to  a  fragile  coal,  or  to 
other  similar  conditions  requiring  ample  support.  The  stall  system  is 
particularly  useful  in  deep  seams  where  the  roof  pressure  is  great.  The 
stalls  are  usually  opened  narrow  and  widened  inside  to  furnish  a  breast 
which  varies,  according  to  conditions  of  roof,  floor,  coal,  depth,  etc., 
from  4  to  6  yd.  wide  in  the  " single-stall"  method.  The  pillars  between 
the  stalls  are  usually  about  the  width  of  the  breasts. 

Fig.  132  shows  the  method  by  single  stall  and  Fig.  133  that  by  double 
stall.  The  former  is  more  applicable  to  flat  seams  or  seams  of  small 
inclination;  while  the  latter  is  used  on  steep  pitches.  The  single-stall 
method  affords  but  one  road  to  a  breast;  and,  hence,  does  not  permit  of 


264  MINING    WITHOUT   TIMBER 

the  concentration  of  men  possible  in  double  stalls  where  there  are  two 
roads  to  each  breast.  In  the  double  stalls  the  breasts  are  wider,  ranging 
from  12  to  15  yd.;  while  the  pillars  sometimes  reach  a  width  of  30  yd. 

Panel  System. — It  is  advisable  to  mine  in  panels:  1,  When*  the  seam 
contains  much  gas,  making  it  essential  that  the  ventilation  of  the  entire 
mine  be  under  absolute  control;  2,  when  the  coal  is  readily  affected  by 
the  air,  and  disintegrates  with  long  standing;  3,  When  the  roof  pressure 
or  the  conditions  of  the  roof  are  such  as  to  require  extreme  caution  to 
prevent  squeeze  or  creep.  The  panels  are  formed  by  driving  entries  and 
cross  entries  so  as  to  intersect  each  other  at  regular  intervals  of,  usually, 
about  100  yd.  The  entire  field  is  thus  ultimately  divided  into  separate 
squares  or  panels,  each  of  which  has  practically  its  own  system  of  ven- 
tilation. Each  alternate  haulway  may  be  made  an  intake  to  supply  air 
to  one  tier  of  panels,  while  the  next  succeeding  passageway  may  be  used 
as  the  return  to  conduct  the  air  from  each  panel  to  the  foot  of  the  upcast. 
If  more  air  than  usual  is  needed  in  any  one  panel,  it  can  be  obtained  at 
once  by  enlarging  the  opening  in  the  regulator  which  controls  the  air 
for  that  section.  Irr  case  of  an  explosion  in  any  one  panel,  it  is  not 
usually  communicated  to  the  other  panels.  Extractions  can  be  com- 
menced as  soon  as  a  panel  is  formed;  and  usually  consist  in  driving  a 
heading  across  the  panel  and  opening  the  coal  by  single  o-r  double  rooms 
or  stalls.  Next,  the  room  pillars  are  carefully  drawn  and  the  roof  inside 
the  peripheral  pillar  of  the  panel  is  allowed  to  fall.  A  high  extraction 
of  coal  can  thus  be  safely  secured  with  a  small  loss  of  timber. 

EXAMPLE  55. — NELMS'  RETREATING  SYSTEM 

Room  and  Pillar  Layout  for  Flat  Coal  Seams. — This  method  insures 
the  operator  a  greater  amount  of  coal  than  when  the  seam  is  worked 
advancing  on  the  room  and  pillar  system.  Since  mining  men  in  the 
United  States  now  recognize  that  our  supply  of  fuel  is  exhaustible,  it 
certainly  behooves  all  operators  to  mine  every  ton  of  coal  possible. 

In  this  retreating  system,  the  main  entries  are  driven  50-ft.  centers 
with  cross-cuts  every  100  ft.  The  middle  entry,  in  the  three-entry 
system,  is  used  for  the  haulage  road,  being  also  a  main  intake  airway. 
After  turning  a  pair  of  butt  entries  off  the  main,  the  second  crosscut, 
200  ft.  from  the  last  butt  entry,  should  be  a  45-deg.  chute  for  motor 
haulage.  The  dotted  lines  on  the  main  entry  at  the  bottom  of  the  butts 
show  the  position  of  the  "  parting."  The  motor,  hauling  25  1  1/2-ton 
cars  comes  in  the  middle  main  entry,  swinging  its  trip  of  empties  in  the 
chute,  the  motor  running  up  the  straight  where  the  drivers  have  stocked 
their  loaded  coal.  The  motor  can  then  pull  its  loaded  trip  outside  and 
the  drivers  proceed  to  distribute  their  cars,  two  drivers  going  in  each 
butt  entry.  The  drivers  make  two  trips,  while  the  motor  makes  one. 


PILLAR   SYSTEMS   FOR   SEAMS 


265 


The  butt  entries  are  driven  on  a  90-deg.  angle  from  the  main  entries, 
and  at  a  distance  of  1400  ft.,  they  intersect  a  set  of  three-face  entries 
running  parallel  to  the  main  entries.  The  butts  are  driven  50-ft.  centers, 
with  crosscuts  every  100  ft.  This  system  of  turning  butts  off  the  mains 
is  an  ideal  one  for  haulage  and  ventilation.  Instead  of  driving  rooms 
off  the  butts  beginning  near  the  main 
entry,  the  rooms  are  started  from  the 
face-entry  side  and  all  coal  is  worked 
toward  the  main  entries. 

Usually  60  ft.  of  solid  coal  is  left  to 
protect  the  face  entries,  and  60  ft.  to 
also  protect  the  mains.  The  rooms 
are  started  four  at  a  time,  and  as  soon 
as  the  first  four  have  been  driven  50 
ft.,  the  next  four  are  started  on  both 
butts.  The  rooms  are  all  driven  on 
sights  90  deg.  off  the  butt  entry  and 
driven  25  ft.  wide  for  a  distance  of 
240  ft.,  there  being  a  15-ft.  pillar  left 
in  each  room.  The  crosscuts  in  the 
rooms  are  from  80  to  100  ft.  apart,  and 
should  be  " staggered"  across  the 
different  rooms  so  as  not  to  make  a 
weak  place  in  the  roof  by  having  the 
breaks  all  opposite. 


FOUR  PAIRS  OF  BUTT  ENTRIES  WILL 
PRODUCE  1200  TONS  OF  COAL  DAILY 

After  driving  the  rooms  the  full 
distance,  they  should  be  cut  over  to 
the  next  room  by  the  mining  ma- 
chine, the  cut  being  20  ft.  wide.  The 
great  advantage  to  be  gained  in  this 
system  is  the  method  of  not  having 
work  scattered  all  over  a  mining- 
territory.  Four  pairs  of  butt  entries,  thus  mined,  will  produce  1200 
cars  of  coal  each  working  day. 

In  Fig.  134,  there  are  32  " machine"  rooms  (Nos.  13  to  28  on  each 
side)  working  on  the  pair  of  butts,  and  requiring  32  men  (loaders),  two 
men  having  two  rooms  and  working  them  together.  It  is  the  general 
practice  to  clean  up  one  room  at  a  time  and  so  always  have  coal  to  load 
in  one  room  or  the  other.  Each  machine  loader  receives  six  cars,  thereby 
producing  192  cars  per  day. 


FIG.  134. — Plan  of  N elms'  Retreating  system. 


266  MINING    WITHOUT    TIMBER 

There  are  10  pillars  being  robbed  (Nos.  6  to  10  on  each  side),  and 
these  produce  30  cars  of  coal,  as  the  pillars  are  worked  by  one  man.  In 
some  places  two  men  work  the  pillars.  The  "turn"  in  coal  mines  is 
such  that  a  machine  loader  receives  two  cars  to  the  pick  miner's  one, 
thereby  equalling  each  other's  wages,  as  pick  costs  about  twice  as  much 
as  machine  coal.  The  chain  pillar  and  stump  will  produce  12  cars, 
with  four  men  working,  and  the  two  butts  yield  234  cars  per  day. 

The  engineer  can  advance  the  work  in  such  a  standard  way  that  his 
machine  coal  will  always  total  to  the  proper  amount.  A  mine  foreman 
should  find  this  an  easy  way  to  keep  his  men  standardized,  the  machine 
loaders  always  having  machine  places  and  the  pick  men  pick  places, 
thereby  increasing  the  safety  factor  of  his  mine,  as  his  machine  men 
would  never  have  to  do  pick  work. 

The  ventilation  shown  by  the  arrow  heads  is  the  most  practical  to 
use;  the  splits  are  shown  and  also  the  overcast  at  the  bottom  of  the  butt 
entry,  there  being  a  regulator  in  this  overcast.  The  motor  road  is  clear 
of  doors  on  the  main  entries.  The  arrangement  of  chutes  on  the  left 
side  would  be  slightly  different. 

EXAMPLE  56. — NELMS'  ADVANCING-RETREATING  SYSTEM 

Room  and  Pillar  Layout  for  Flat  Coal  Seams. — The  layout  for  the 
advancing-retreating  system  is  as  follows  in  Fig.  135 :  Three  face  entries, 
on  50-ft.  centers,  are  driven  parallel  to  the  main  entries  at  1400-ft. 
intervals.  The  sectional  area  of  the  face  entries  is  kept  as  nearly  as 
possible  to  a  60-ft.  standard  in  a  5-ft.  seam. 

It  is  advisable  where  possible  to  use  all  three  entries  for  intake  air- 
ways; then  the  middle  entry  can  be  used  for  a  haulage  road  and  should 
be  confined  to  itself  and  not  enter  into  the  ventilation  at  all.  No.  1 
room  on  the  butt  entry  can  be  driven  16  ft.  wide  and  used  as  a  return 
airway.  When  the  No.  1  room  is  maintained  for  an  airway,  it  should  be 
widened  toward  the  face  or  main  entry  and  be  driven  on  70-ft.  centers 
with  the  face  entry.  A  30-ft.  pillar  of  solid  coal  should  be  left  between 
No.  1  and  No.  2  rooms;  No.  1  rib  can  then  be  easily  extracted  when 
robbing  is  commenced  on  this  butt  entry. 

The  gob  in  No.  I  room  should  be  kept  as  low  as  possible  and  if  easily 
handled,  it  should  all  be  loaded  and  dumped  outside;  the  result  is  a 
return  airway,  16x6  ft.  =96  sq.  ft.,  which  is  usually  large  enough. 

Butt  entries  should  be  turned  every  450  ft.  A  chute  is  driven  on  a 
60-deg.  angle,  from  the  middle  main  entry  to  the  outside  main  for  haulage. 
The  butts  are  turned  on  a  '90-deg.  angle,  and  are  driven  on  50-ft.  centers 
for  a  distance  of  1400  ft.  A  60-deg.  chute  should  connect  the  butt 
entries  at  25  ft.  from  the  center  of  the  outside  main,  for  haulage  from 
he  butt  entry.  It  is  bad  practice  cutting  corners  off  break-throughs. 


PILLAR   SYSTEMS    FOR   SEAMS 


267 


Each  butt  entry  should  maintain  a  sectional  area  of  about  50  sq.  ft. 
and  be  driven  perfectly  straight  so  as  to  overcome  the  troubles  of  track 
laying,  cars  jumping  track,  etc.  Entry  sights  should  never  be  more  than 
180  ft.  apart,  so  as  to  allow  the  mine  foreman  a  good  chance  to  keep  his 
sights  well  up.  In  providing  ventilation,  two  pairs  of  butts  on  one  split 
are  suitable.  The  rooms  should  be  turned  90  deg.  off  butts  and  driven 
as  shown  in  plan. 


UL 


Faces  of  Rooms  to  be  all  Eveu  ou  240 'Line 
FIG.  135. — Plan  of  Nelms'  Advancing-retreating  system. 


GENERAL  LAYOUT 

No.  1  room  should  be  started  as  soon  as  possible,  then  No.  2  and  so  on 
from  the  advancing  butt  entry.  When  No.  2  room  is  finished  working 
on  the  face,  No.  14  should  just  be  started;  when  No.  2  rib  is  out,  No.  14 
room  should  be  just  finished  and  No.  27  room  just  starting.  The  ribs 
must  be  extracted  as  soon  as  each  room  is  finished,  no  matter  whether 
the  next  room  on  the  advance  side  is  finished  or  not;  the  rib  is  started 
while  the  next  room  is  still  30  ft.  from  being  finished.  When  No.  32  is 
finished  working  on  the  face,  No.  19  rib  is  just  finished,  also  No.  1  room 
on  the  retreating  butt  is  about  finished  and  No.  13  room  just  started. 
As  soon  as  No.  2  rib  on  the  retreating  entry  is  finished,  it  is  advisable  to 
start  extracting  immediately  the  butt-entry  stumps  and  chain  pillar, 
bringing  everything  along  with  the  retreating  butt  and  closing  entry  in 
tight,  knocking  out  the  brattice  in  each  succeeding  break-through  for 
ventilation. 


268 


MINING    WITHOUT   TIMBER 


This  method  allows  the  rib  men  and  the  machine  loaders  to  be  always 
separate,  the  workings  are  confined  to  the  smallest  space  possible  for  a 
large  tonnage,  and  ventilation  is  easy.  The  mine  will  not  be  dotted 
with  old  abandoned  workings  if  the  method  is  consistently  maintained. 

When  a  set  of  butts  are  thus  worked  out  they  are  off  the  operator  7s 
hands.  There  is  never  any  danger  of  a  squeeze  as  every  movement  of 
the  rock  runs  up  against  solid  coal,  and  for  this  reason  it  is  impossible  to 
have  a  squeeze  swing  across  a  set  of  butts  to  another  set,  as  it  generally  does 
where  both  entries  are  worked  advancing.  Five  pairs  of  butts  developed 
on  this  plan  can  produce  from  1000  to  1500  cars  of  coal  a  day. 

Small  mule  "partings"  should  be  at  the  bottom  of  each  pair  of  butt 
entries  and  the  distance  for  mule  haulage  will  then  be  at  a  minimum. 
The  coal  from  these  partings  can  be  gathered  by  a  6-  or  8-ton  locomotive, 
and  delivered  to  a  longer  parting  whence  a  larger  locomotive  can  take  it 
outside. 


FIG.  136. — Plan  for  pillar-drawing,  Connellsville. 


EXAMPLE  57. — CONNELLSVILLE  COKE  DISTRICT,  WESTERN  PA. 
(See  also  Example  58.) 

Retreating  System  on  Thick,  Flat  Pittsburg  Seam. — The  plan  of  mining 
in  the  Connellsville  region  has  grown  from  the  primitive  methods,  suitable 
to  the  favorable  mining  conditions  when  operations  were  first  started,  to 
the  scientific  methods  which  became  necessary  as  the  cover  increased, 
and  when  most  of  the  difficulties  that  are  likely  to  be  met  in  deep  mining- 
were  encountered  and  overcome.  The  conditions  here  are  that  all  the 
coal  is  coked  so  that  fine  coal  is  an  ad  vantage;  no  machines  are  used,  but 
the  coal  is  dug  with  pick;  the  seam  averages  7  1/2  ft.  of  clean  coal  mined; 
the  roof  is  friable  and  some  coal  is  left  in  the  top  to  support  it;  the  over- 
lying stratum  is  generally  6  to  10  ft.  of  slaty  coal  with  sandstone  above. 

Where  the  acreage  owned  or  assigned  to  each  mine  is  too  large  to 


PILLAR   SYSTEMS    FOR    SEAMS 


269 


admit  of  going  to  the  extreme  boundary  before  starting  to  draw  the 
ribs,  it  is  customary  to  divide  the  field  into  panels  1000  or  1500  ft.  square, 
by  face  headings  driven  off  the  main  butt  headings  as  shown  in  Fig.  136. 
The  coal  is  first  removed  from  the  extreme  side  of  each  panel  away  from 
the  main  butt  headings,  a  diagonal  break  line  is  established,  as  shown, 
and  the  coal  withdrawn,  retreating  toward  the  near  corner,  keeping  the 
break  line  straight,  and  the  coal  between  where  the  drawing  is  being 
carried  on  and  the  main  butt  headings  as  nearly  solid  as  possible,  the 
butt  headings  and  rooms  being  drawn  only  fast  enough  to  open  up  the 
coal  for  the  drawings.  The  break  line  of  roof  is  kept  as  nearly  perpen- 
dicular as  practical  to  direction  of  rooms.  The  recent  tendency  in  the 
large  mines  is  not  to  start  pillar-drawing  till  the  boundary  is  reached. 

*%$&&'&&£ 

c>>4,   .  .'-j  ..:  . .- . .  ."•f. 

d     a| 

'ss' 


FIG.  137. — Pillar-drawing  with  thin  cover,  Connellsville. 


The  face  of  the  coal  in  this  region  is  well  defined  on  a  line  running 
N  17°  E.  Where  the  grades  are  not  too  great  all  headings  are  driven 
square  on  the  face  or  on  the  butt,  and  the  rooms  always  on  the  face  and 
only  to  the  rise.  The  rooms  are  driven  10  or  12  ft.  wide  and  a  line  of 
posts  set  as  the  room  advances,  as  shown  in  Fig.  137,  the  posts  being  set 
about  4  or  5  ft.  apart.  A  track  of  wooden  or  steel -rails  is  laid  by  the 
miner  close  to  the  upper  rib.  The  width  of  the  pillar,  which  varies  from 
10  to  70  ft.,  is  governed  by  the  softness  of  the  bottom  and  the  thickness  of 
the  overlying  strata. 

There  is  some  variation  in  the  method  of  drawing  the  individual 
ribs  but  the  principle  is  the  same.  On  account  of  the  nature  of  the 
roof,  short  falls  are  necessary,  two  or  three  being  made  before  the  over- 
lying rock  is  broken.  When  the  rock  breaks  it  will  crush  posts  so  that  it 


270 


MINING    WITHOUT    TIMBER 


is  necessary  to  break  the  roof  against  the  end  of  the  ribs  and  not  over  posts. 
If  care  is  taken  the  digger  knows  when  to  expect  a  fall  and  very  few 
posts  need  be  lost. 

Fig.  137  shows  the  method  in  use  when  the  overlying  strata  are  under 
200  ft.  thick  and  where  a  curved  track  and  track  along  the  face  are  not 
put  in.  A  slab  a  is  taken  off  the  right-hand  rib,  the  whole  of  the  left- 
hand  rib  is  taken  out  excepting  enough  only  left  in  to  keep  the  gob  from 
mixing  with  the  coal  as  shown  at  6.  A  small  shoulder  c  is  left  in  the.  far 
corner  of  the  rib  to  take  the  brunt  of  the  weight.  A  row  of  posts  /  is  set 
from  this  shoulder  across  the  room  to  preserve  a  working  face.  The 
posts  e  in  the  back  of  this  row  and  between  it  and  the  previous  fall  or 
break  are  drawn  and  any  which  cannot  be  drawn  are  cut  so  as  to  get  a 


FIG.  138. — Pillar-drawing  under  thick  cover,  Connellsville. 

», 

good  fall.  Two  men  work  in  each  room.  Room  40  shows  the  situation 
in  the  last  room  of  a  tier  of  rooms  along  a  butt  heading.  Slabs  are  being 
taken  off  each  side  pillar  while  the  men  are  protected  by  the  row  of  posts  d. 

Room  39  shows  the  condition  of  a  room  just  before  producing  a  fall 
by  withdrawing  the  row  of  posts  d  and  separate  posts  e  between  the  rows 
d  and/.  The  row  of  posts/  and  the  stump  of  coal  c  protect  the  face  dur- 
ing the  fall.  The  stump  c  is  next  removed  excepting  for  the  small  slab 
6  and  the  work  proceeds  as  shown  at  the  face  of  room  40. 

Fig.  138  shows  a  method  of  drawing  pillars  where  there  is  over  200  ft. 
of  surface  above  the  coal  and  where  a  curve  is  used  to  run  the  track  along 
the  face.  The  rooms  are  driven  12  ft.  wide  while  the  pillars  between 
the  rooms  vary  from  34  ft.  up  to  60  ft.  All  of  the  rooms  are  driven  on 
sights  so  that  the  pillars  may  be  of  uniform  thickness.  After  the  last 
room  on  the  heading  is  driven  the  required  length,  which  is  about  300  ft., 
the  pillar  is  cut  across  at  the  face  of  the  room  and  2Q  or  30  ft.  removed 


PILLAR    SYSTEMS    FOR    SEAMS 


271 


before  drawing  the  posts  and  getting  a  fall  of  the  roof.  The  usual 
method  is  then  for  two  men  to  work  on  each  pillar,  while  one  man  cuts 
back  in  the  center  of  the  pillar  on  the  face  of  the  coal  as  far  as  he  can 
conveniently  shovel  as  shown  at  a,  room  27,  the  mine  car  being  on  track 
6,  the  other  man  is  drawing  stump  c  and  shoveling  into  the  same  car. 

When  c  is  all  removed  a  fall  is  made  and  the  situation  is  similar  to 
that  shown  in  rooms  30  and  26,  the  curved  track  having  been  removed 
to  one  side  and  a  straight  one  substituted.  Now  one  man  cuts  into  the 
side  of  the  pillar  8  ft.  from  the  end  at  d  as  shown  while  the  other  is  re- 
moving the  stump  e.  When  this  is  accomplished  a  fall  is  made  and  the 
curve  put  in,  the  conditions  being  then  as  shown  in  room  29;  the  two 
men  then  continue  to  cut  over  toward  the  gob  in  the  next  room  as  shown 
in  room  29,  the  curved  track  having  meanwhile  been  put  along  the  face. 
Room  28  shows  the  situation  when  the  two  men  have  driven  this  cut 
through  to  the  gob  in  the  adjoining  room.  Room  27  illustrates  the  next 


Rib  cut  across  Rib  ready  for  first  fall.       Rib  ready  for  last  fall. 

Tlie  Enftineerma  if  Mining  Journal 

FIG.  139. — Second  method — pillar-drawing  under  thick  cover,  Connellsville. 

step.  While  one  man  works  on  the  pillar  c  in  the  far  corner  of  the  room 
the  other  starts  the  cut  a  back  into  the  face  as  shown.  The  curved  track 
b  is  then  lifted  and  a  straight  track  put  in  as  shown  in  room  26  in  order  to 
get  out  the  pillar  e.  While  one  man  is  removing  this  pillar  the  other 
one  starts  to  cut  into  the  rib  as  shown  at  d,  room  26. 

When  a  fall  is  to  be  made  posts  are  set  18  in.  apart,  as  shown  in  room 
26  across  the  end  of  the  room  and  along  the  end  of  the  hole  into  the 
pillar.  All  of  the  other  posts  beyond  the  break  rows  are  drawn.  The 
curved  track  is  laid  into  the  new  cut  in  the  side  of  the  pillar  at  d  and  by 
the  next  morning  the  roof  has  fallen.  In  places  where  the  rib,  is  wider 
than  shown  on  the  plan  a  couple  of  falls  can  be  made  in  the  width  of  the 
pillars  by  placing  break  rows  of  props  similar  to  those  already  described. 

A  third  pillar-robbing  method,  for  a  heavy  cover,  is  shown  in  Fig.  139. 
It  cross-cuts  the  pillar  for  a  track  and  uses  the  same  curve  as  the  previous 
method,  but  instead  of  cutting  up  the  resulting  pillar-slab  into  blocks, 
it  begins  at  the  gob  and  withdraws  the  slab  gradually  roomward,  mean- 
while recovering  most  of  the  props  in  the  figure  that  were  put  in  to 
protect  the  excavation  of  coal.  A  ratchet-chain  puller  for  props  is  used 
where  necessary.  The  use  of  this  method  at  the  Continental  No.  1  mine 


272  MINING    WITHOUT   TIMBER 

of  the  Frick  Coal  Company  gives  a  recovery  of  about  90  per  cent,  of  the 
coal.  The  losses  arise  from  a  6-in.  coal  layer,  impure  with  sulphur,  left 
on  the  floor;  a  coal  layer,  of  4  in.  in  the  rooms  and  9  in.  in  the  entries, 
left  on  the  roof;  and  some  occasional  stumps  lost  in  robbing  the  pillars. 

The  foremen  of  the  district  are  guided  by  the  following  10  rules  in 
extracting  pillars.  (1)  Pillar  robbing  must  not  be  stopped  or  diverted 
from  the  line  of  fracture  without  the  consent  of  the  chief  engineer.  (2) 
Robbing  must  proceed  from  the  new  toward  the  older  gob  to  prevent 
uncalculable  pressure  on  the  working  face.  (3)  Ribs  must  be  robbed 
within  one  month  of  driving  rooms.  (4)  Room  centers  must  be  at  the 
prescribed  distance  apart.  (5)  In  robbing  entry-pillars,  a  length  of 
only  2  room-widths  must  be  attacked  at  once  along  the  line  of  fracture. 
(6)  Water  ditches  must  be  made  for  entry-drainage  and  especial  care 
must  be  taken  on  soft  bottoms.  (7)  A  200-ft.  pillar  must  be  left  on 
each  side  of  the  main  or  flat  entry  during  its  life.  (8)  A  wide  barrier 
pillar  must  be  left  and  care  must  be  taken  in  approaching  a  neighbor's 
boundary.  (9)  Before  permitting  a  fall  of  the  roof,  all  timber  must  be 
drawn  and  a  passage  left  for  the  escape  of  the  miners.  (10)  A  miner 
should  keep  the  pillar  he  is  drawing  between  himself  and  the  gob 
instead  of  working  between  gob  and  pillar. 

EXAMPLE  58. — PITTSBURG  BITUMINOUS  DISTRICT,  WESTERN  PA. 
(See  also  Example  57.) 

Advancing-retreating  or  Retreating  System  in  Panels  on  Thick  Flat 
Pittsburg  Seam. — In  those  mines  of  western  Pennsylvania,  extracting 
the  thick  Pittsburg  coal  seam  for  shipment  to  market,  the  mining  layout 
is  different  from  that  of  the  coking  district  of  Example  57.  Since  the 
policy  of  the  market-coal  mines  is  to  obtain  as  much  lump  coal  as  possible, 
the  bulk  of  the  coal  is  obtained  from  the  rooms,  for  pillar  coal  is  bound  to 
be  more  or  less  crushed.  This  policy  requires  wide  rooms  and  narrow 
pillars  and  results  in  a  lesser  total  recovery  of  coal,  but  as  an  offset,  more 
coal  can  be  won  by  machine  cutters  which  work  advantageously  in  this 
thick  seam.  By  the  deep  and  fast  undercutting  possible  with  machines, 
blasting,  with  its  ensuing  slack,  is  at  a  minimum;  and  progress  is  rapid 
enough  to  preserve  the  coal  faces  from  long  exposure  to  the  atmosphere 
and  to  allow  of  systematic  timbering  and  an  even  subsidence  of  the  roof. 

The  Monongahela  River  Cons.  Coal  and  Coke  Company  is  a  very 
large  producer,  and  a  composite  drawing  of  its  method  of  working  is 
shown  in  Fig.  140.  Above  the  upper  "butts"  or  butt  entries,  the  rooms 
have  not  advanced  far  from  the  "Face  Entries."  On  the  middle  butts, 
the  rooms  have  reached  the  end  of  the  upper  panel  and  pillar-drawing 
has  advanced  halfway.  On  the  lower  butts,  the  lower  panel  is  being 
worked  by  retreating  from  the  panel-end,  the  rooms  are  nearly  completed, 


PILLAR   SYSTEMS    FOR   SEAMS 


273 


and  the  line  of  roof-fracture,  across  both  upper  and  lower  panels,  is 
following  the  pillar^drawing  and  is  not  far  behind  the  finished  rooms. 

In  the  advancing  system,  the  panels  off  the  upper  butt  entry  would 
be  attacked  first,  and  the  rooms  of  this  advancing  panel  would  end  in  the 
old  gob,  while  the  rooms  of  the  lower  or  retreating  panel  would  end 
against  the  solid.  As  shown  in  the  figure,  the  room-centers  are  39  ft. 


Return  Current 
Brick  Stoppings    =• 
Doors  /"• 

Overcasts 


FIG.  140. — First  layout  (Monongahela  colleries)  at  Pittsburg,  Pa. 

apart,  of  which  space  the  pillars  occupy  15ft.  It  will  be  noticed  that  the 
room-stumps  of  each  upper  panel  are  left  undisturbed  on  the  advance  so 
as  to  protect  the  return  airway,  but  when  the  pillars  of  the  lower  panel 
are  being  drawn,  the  upper  stumps  are  also  pulled,  as  the  receding  line  of 
roof-fracture  passes  them,  along  with  the  butt-entry  pillars.  The  last 
pillars,  however,  must  be  left  undisturbed  in  the  advancing  system  along 

18 


274 


MINING    WITHOUT    TIMBER 


their  whole  length  until  all  the  adjoining  coal  has  been  exhausted  up  to 
the  boundary.  The  use  of  four  main  entries,  by  this  method,  allows  the 
two  outside  gangways  to  be  return-airways  and  the  two  intake-airways 
to  be  on  the  inside  and  thus  gives  an  ample  main-airway  area  and  a  min- 
imum interference  with  transport.  The  room- work  is  in  the  fresh  air  and 
pillar-drawing  is  on  the  return-air  side  of  it.  The  room-track  is  always 
laid  along  the  straight  rib,  and  in  many  mines  the  refuse  between  the 
track  and  the  other  rib  fills  the  room  nearly  roof  high. 


*n 


Fia.  141. — Second  layout  for  large  output,  Pittsburg. 


Fig.  141  shows  a  second  layout  for  large  output,  used  in  the  Pittsburg 
seam,  with  six  main  entries.  There  are  three  face  entries,  nominally,  but 
four  actually,  as  the  nearest  room  on  the  butt  is  advanced  along  with 
them  so  as  to  give  an  additional  airway.  As  shown,  the  rooms  are  only 
worked  on  the  outbye  side  of  the  butts,  and  the  first  room  is  started 
from  the  far  end  of  a  panel  and  followed,  at  the  proper  distance  on  the 
retreat,  by  pillar-drawing.  By  starting  work  from  No.  2  and  the  follow- 
ing butts  at  the  proper  time,  it  is  possible  to  keep  the  line  of  roof -fracture 


PILLAR    SYSTEMS   FOR   SEAMS 


275 


of  a  panel  continuous,  for  entry-pillars  and  room-stumps  are  removed 
as  shown  from  the  panel-end  back  to  the  butts. 

The  method  of  Fig:  141  has  permitted  the  extraction  of  70  per  cent. 
of  the  pillars  by  machine  cutters  under  an  average  cover  of  200-ft.  thick- 
ness. For  this  purpose  machine  cross-cuts,  21  ft.  wide,  are  made  in  the 
pillars  so  as  to  leave  for  each  a  stump  only  9  ft.  wide  to  be  removed  by 
hand-pick.  This  cross-cutting  is  shown  by  the  different  cross-hatching 
of  the  figure  which  also  illustrates  the  overcasts  and  brattices  for  ventila- 
tion, and  the  chutes,  etc.,  for  transport. 

A  third  method  of  attack  by  which  one  company  mines  over  2,500,000 
tons  yearly  is  shown  in  Fig.  142.  Here  the  room  pillars,  after  the  room 


FIG.  142. — Third  layout,  with  tapering  pillars,  Pittsburg. 

has  advanced  100  ft.,  or  to  the  first  break-through,  are  gradually  tapered 
off  to  a  point  at  the  room-end.  This  causes  the  roof  to  fall  along  the 
tapered  parts  of  the  pillars  and  the  latter  are  lost,  but  much  of  the  thicker 
pillar  near  the  room-neck  can  be  recovered  by  subsequent  careful  pick- 
work.  This  method  gets  nearly  all  the  coal  by  room-work,  and  a  total 
recovery  of  90  per  cent,  is  claimed  by  its  advocates.  It  is  more  danger- 
ousrhowever,  than  the  two  previous  systems,  requires  more  timber,  and 
squeezes  are  more  liable  to  occur. 

Where  this  high  Pittsburg  seam  is  dirty,  so  that  much  gob  must  be 
stowed  along  the  rib  on  the  advance,  it  is  customary  on  drawing  the 
pillars  to  leave  a  vertical  shell  of  from  12  to  18  in.  of  coal  next  to  the 
gob  to  prevent  any  pollution  of  the  broken  coal. 


CHAPTER  XXI 

FLUSHING  SYSTEM  FOR  FILLING  SEAMS  AND  RECOVERING 

PILLARS 

EXAMPLE  59. — ANTHRACITE  DISTRICT,  EASTERN  PA. 
(See  also  Examples  5,  51  and  59.) 

Parallel  Seams  of  Various  Thickness  and  Dip  Filled  with  Refuse  from 
Breakers  and  Dumps. — The  flushing  system  was  first  developed  in  1891 
at  the  Dodson  mine  near  Wilkesbarre,  Pa.;  and  was  later  copied  and 
extensively  used  in  many  German  collieries.  Three  conditions  made 
flushing  a  valuable  innovation  in  the  Pennsylvania  anthracite  region, 
namely,  the  numerous  large  dumps  of  waste  available  for  filling,  the 
parallel  and  superincumbent  seams  to  be  extracted,  and  the  overlay  of 
much  workable  coal  by  townsites.  The  gravity  of  the  urban  situation 
is  evidenced  by  the  report  of  April,  1911,  made  by  the  Scranton  Com- 
mission. This  report  states  that  a  large  part  of  Scranton  is  already 
undermined  and  that  for  the  stability  of  the  present  dangerous  area  of 
15  per  cent,  of  the  city  and  for  the  recovery  of  the  coal  pillars  from  the 
balance  the  flushing  system  is  the  only  remedy. 

The  following  description  in  based  on  the  author's  visits  to  mines  of 
the  following  coal  companies:  Philadelphia  and  Reading;  Delaware 
and  Hudson;  Delaware,  Lackawanna  and  Western;  Plymouth;  and 
Lehigh  Valley. 

In  mining  the  flat  seams  to  the  north  of  Wilkesbarre  by  the  pillar 
system  of  Fig.  130  much  of  the  waste  broken  with  the  coal  can  be  left 
in  the  rooms;  but  in  the  seams  of  the  southern  districts  where  mining  is 
done  by  "  overhand  stoping  with  shrinkage  and  chutes,"  as  in  Figs.  132 
and  133,  all  the  waste  has  to  be  hoisted.  The  crude  coal  reaching  the 
surface  is  a  mixture  of  pure  coal,  " slate,"  " slate-coal",  and" bone." 
The  "slate"  corresponds  to  the  shale  and  clay  of  the  partings  and  beds 
of  the  bituminous  regions,  the  "slate-coal"  consists  of  lumps,  part  pure 
coal  and  part  slate,  and  the  "bone"  is  a  coal  containing  too  little  carbon 
(present  limit  60  per  cent.)  to  be  marketable.  All  crude  coal  is  put 
through  a  dressing  mill  or  "  breaker"  in  which  impure  pieces  are  broken 
sufficiently  to  detach  the  slate  and  bone  from  the  pure  coal,  so  that  all  the 
latter  may  be  screened  for  separation  into  commercial  sizes  and  the 
former,  along  with  the  "culm"  or  coal  dust,  be  sent  to  the  waste  dump 
or  mine  stopes. 

276 


FLUSHING   SYSTEMS    FOR   FILLING    SEAMS 


277 


278  MINING    WITHOUT    TIMBER 

The  limit  of  size  between  fine  coal  and  unmarketable  "culm77  has  so 
decreased  in  recent  years  that  now  all  the  fine  sizes  of  the  breakers,  as 
well  as  many  old  waste  dumps,  are  being  washed  over  shaking  screens 
in  special  mills  called  "washeries,"  for  their  content  of  fine  coal  of 
commercial  value.  The  present  upper  limit  for  "culm7'  is  a  diameter 
varying  from  5/64  to  3/16  in.,  but  some  independent  operators  use 
also  some  larger  sizes  for  flushing.  This  rejected  fine  coal,  as  mixed  with 
the  slate  and  bone  tailing  from  the  breaker  and  the  ashes  from  the  boiler 
plant,  forms  "  slush, "  the  chief  material  now  used  in  filling  the  mine  stopes 
by  the  flushing  system.  At  some  mines,  the  larger  pieces  of  bone  are 
saved  on  a  special  dump  as  of  possible  future  value.  Fig.  143  shows 
the  Dodson  colliery  at  Plymouth,  Pa.,  with  the  waste  dump  at  A,  the 
breaker  at  B,  and  the  washery  at  C. 

In  digging  an  old  waste  dump  for  passage  through  a  washery,  in 
order  to  separate  the  marketable  coal  before  flushing,  a  system  of  chain 
conveyors  as  at  D  and  H,  Fig.  143,  is  used.  The  usual  conveyor  has  a 
single  chain  and  drags  its  steel  plate  scrapers,  18  in.  long,  12  in.  high, 
and  set  3  ft.  apart,  in  a  trapezoidal  trough  made  by  lapping  the  ends  of 
3-ft.  lengths  of  steel  or  cast-iron  plate.  The  maximum  length  of  a  single 
conveyor  trough  is  about  500  ft.  It  is  supported  within  square  wooden 
frames,  E,  set  8  ft.  apart,  and  built  of  4x6-in.  pieces.  Near  the  top  of 
the  frames  E,  run  two  25-lb.  steel  rails  to  support  the  scrapers  on  their 
return  trip. 

Each  conveyor  is  run  by  an  independent  steam  engine,  as  at  F,  con- 
nected by  gearing  to  its  head  end,  and  its  capacity  of  100  to  200  tons  of 
dry  material  per  hour  is  fed  in,  anywhere  along  the  trough,  by  hand 
shovels  or  by  hydraulicing  with  hose.  Obstacles  between  the  dump 
and  washery  are  passed  by  using  several  conveyors,  set  at  an  angle,  of 
which  only  the  conveyor  at  the  feed  end  need  be  shifted  as  the  dump 
dwindles.  The  driving  engine  is  set  on  a  timber  frame  so  that  it  can  be 
easily  pushed  into  line,  by  screw  jacks,  when  the  conveyor  is  moved 
over  by  levers;  both  engine  and  conveyor  are  elevated  on  rollers  before 
shifting.  This  is  done  by  the  regular  attendants  who  consist  of  two  men 
for  feeding  and  one  man  at  each  driving  engine. 

So  much  fine  marketable  coal  can  now  be  saved  from  the  present 
breaker-tailing  and  from  the  old  "culm7'  dumps  that  the  final  rejected 
waste  can  fill  only  a  fraction  of  the  space  left  above  the  gob  in  the  under- 
ground rooms.  The  huge  dumps  which  formed  such  a  prominent 
feature  of  the  landscape,  as  late  as  the  early  nineties,  are  rapidly  dis- 
appearing and  filling  is  now  being  won  even  from  river  beds. 

Thus  the  Plymouth  Coal  Co.  has  a  plant  to  bring  sand  and  fine  mine 
waste,  now  settled  at  the  bottom  of  the  Susquehanna  river,  to  its  Dodson 
mine  No.  12.  A  suction  pump  on  a  barge  located  across  the  river  from 
the  Dodson  breaker  delivers  into  a  pipe  which  crosses  the  river  on 


FLUSHING    SYSTEMS   FOR   FILLING    SEAMS  279 

barges  and  discharges  into  an  elevator  which  lifts  the  material  to  the 
flushing  flume  for  the  mine. 

As  no  pieces  larger  than  1-in.  dia.  and  few  over  1/4-in.  dia.  are  used 
in  flushing,  the  coarser  pieces  of  bone  and  slate,  from  breaker  or  dump, 
are  passed  through  a  pulverizer  usually  of  the  Williams'  or  Jeffrey's  type, 
before  reaching  the  flume  G,  Fig.  143,  where  they  mix  with  the  fine 
waste  from  washery  C.  Enough  water  is  put  in  the  flume  to  transport 
the  slush  along  the  flat  pipes  above  the  stopes,  so  the  liquid  slush  carries 
only  about  20  per  cent,  of  solids  by  weight.  The  descent  of  the  slush  is 
through  wrought-iron  pipes,  4  to  6  in.  dia.,  following  either  a  shaft  or  a 
special  bore-hole  into  the  workings,  perhaps  1000  ft.  beneath.  Over  the 
top  of  the  descending  pipe  is  placed  a  funnel  and  a  screen  with  1-in.  holes, 
and  at  each  flushing  station  are  three  gate  valves,  to  regulate  the  flow 
into  the  stopes,  connected  by  electric  signals  with  the  surface.  One  of 
these  station  valves  regulates  the  flow  horizontally,  another  cuts  off  the 
vertical  column  below,  and  by  a  third  the  column  can  be  drained  up  to 
the  nearest  flowing  point  above,  in  case  of  a  stoppage. 

The  pipes  used  for  transport  along  the  levels  are  4  to  6  in.  dia.,  and  of 
either  wrought-iron  or  wood.  In  upgrade  levels  or  where  the  pipe  is  under 
much  pressure,  new  iron  pipes  with  screw  or  flange  couplings  must  be  used, 
but  when  these  are  somewhat  worn  they  are  transferred  to  the  down- 
grade levels.  In  the  latter,  the  iron  pipe  has  standard  couplings  on 
tangents,  but  on  curves  it  has  7-in.  unthreaded  nipples  for  couplings 
slipped  over  the  pipe  ends  and  made  tight  by  wooden  wedges.  These 
wedged  couplings  enable  the  pipe  to  be  rotated,  when  its  bottom  gets 
thin,  so  that  it  can  be  worn  to  a  mere  shell  all  around  before  rejection. 

The  wooden  pipe  is  made  in  Elmira,  N.  Y.,  of  tenoned  maple  staves 
about  2  1/2  in.  thick  which  are  bound  with  spiral  steel  hoops  and  coated 
with  tar.  It  comes  in  2-  to  8-ft.  lengths,  with  male  and  female  ends  for 
slip-jointing  with  cement.  It  can  be  joined  to  cast-iron  fillings  by  in- 
serting special  cast-iron  nipples  in  its  ends,  and  its  own  joints  can  readily 
be  made  to  follow  easy  curves.  It  is  lighter  and  cheaper  than  iron  pipe, 
is  found  to  last  well  on  downgrade  levels,  and  is  preferable  for  use  with 
acid  water.  In  one  mine,  the  iron  pipe  is  used  on  a  2-mile  permanent 
transportation  line  and  the  wooden  pipe  in  the  neighborhood  of  the 
actual  filling. 

Plugged  cast-iron  tees  are  placed  at  100-ft.  intervals  along  all  lines  in 
the  levels,  so  any  obstruction  can  be  easily  located  and  removed.  When 
a  pipe  is  upgrade,  a  special  precaution  is  taken  against  clogging  by  pass- 
ing fresh  water  alone  through  it,  for  15  min.,  before  stopping  the  flow. 
Care  must  be  taken  to  provide  air  escapes  at  high  points  of  the  lines  in 
order  to  avoid  water  hammer. 

The  openings  filled  by  flushing  are  old  rooms  opened  on  the  pillar 
system  of  the  last  chapter.  A  room  on  a  dip  is  easiest  filled,  as  it  requires 


280 


MINING    WITHOUT   TIMBER, 


only  one  dam  or  barrier  at  its  lower  end.  One  disadvantage  of  increasing 
steepness  is  the  greater  strength  of  dam  necessary  to  resist  the  corre- 
spondingly higher  water  head.  In  the  Dorrance  mine  the  old  rooms  had 
been  opened  on  the  rise  from  double  flat  entries  as  in  Fig.  130.  Every 
ten  rooms  along  the  entry  were  separated  by  panel-pillars  following  th.e 
dip.  For  filling,  the  flushing  pipe  was  laid  along  the  airway  above  the 
rooms  and  its  discharge  placed  at  the  head  of  the  central  room  of  a  panel 
of  empty  rooms.  The  latter  had  been  prepared  for  filling  by  erecting 
dams  across  the  necks  at  the  room-bottoms  and  behind  the  break-through 
brattices  of  the  ninth  room 's  pillar,  for  the  last  room  of  the  panel  was  to  be 
left  open  as  an  air  and  manway.  The  brattices  of  the  break-throughs  of 
the  intermediate  rooms  had  been  removed  to  permit  of  a  free  flow  of 
filling  along  the  panel. 

Room  dams  are  made  of  either  stone  or  wood.     The  former  are  thick 
walls  of  roof  slate  laid  up  with  mortar  of  slush  and  straw  in  a  similar 


m 

Roof 

n' 

,    /     iff 

•c  < 

V 

Sec.  a-b 
JIG.  144. — Dam  for  holding  slush,  Eastern  Pa. 


Cross  Sec. 


form  to  the  wall  of  Fig.  145  described  in  the  next  Example.  The  favorite 
dams  are  of  wood  and  a  typical  one  is  shown  in  Fig.  144.  Round  props 
ab,  of  sufficient  size  for  the  expected  strain,  are  covered  on  their  upper 
side  with  2-in.  plank  and  backed,  as  an  extreme  case,  with  stringers  66' 
and  cc'  with  corresponding  angle  braces  bj  and  cd.  If  the  seam-walls  are 
strong,  the  hitches  alone  will  hold  the  props,  so  that  the  pieces  66'  and  bf 
can  be  omitted,  and  in  thin  seams  even  cc'  and  cd  are  left  out. 

When  wetted,  the  seams  between  the  planks  soon  close  up  sufficiently, 
but  the  irregular  spaces  around  the  periphery  mn'b'b  are  caulked  with 
straw  in  one  mine,  and  in  another,  with  a  weak  floor,  the  props  are  set  in 
a  low  concrete  wall,  12  in.  wide.  In  one  seam  of  the  Dorrance  mine  on 
an  18-deg.  slope  with  rooms  300  ft.  long,  the  wooden  dam  of  Fig.  144  is 
strengthened  by  a  dry  wall  of  roof  slate,  3  to  5  ft.  thick,  laid  above  the 
plank  ab.  By  slow  flushing  at  first,  this  dry  wall  gets  packed  solid  and 
keeps  the  plank  from  bulging  under  the  heavy  final  pressure  due  to  a 
vertical  water  head  of  90  ft. 


FLUSHING   SYSTEMS    FOR   FILLING   SEAMS  281 

A  room  in  the  Dodson  mine  in  the  22-ft.  Red  Ash  seam  was  worked 
in  two  slices,  the  first  taking  only  8  ft.  of  coal  from  the  floor.  When 
preparing  for  flushing,  the  upper  14-ft.  slice  of  coal  was  not  taken  down 
over  the  neck  for  24  ft.  from  the  room's  lower  end,  so  that  the  subsequent 
wooden  dam  needed  to  be  only  8  ft.  high.  Holes  are  bored  into  the  plank 
of  the  dams  near  the  top,  if  necessary,  to  let  the  overflow  water  escape, 
but  a  better  arrangement  for  steep  dips  is  a  wooden  drain-launder  bk 
laid  on  the  floor  up  through  the  dam  into  the  room.  The  top  m  of  the 
cover  of  launder  bk  is  kept  a  short  distance  above  the  top  of  the  settled 
slush  at  n  by  adding  new  cover-boards  as  the  filling  rises.  The  overflow 
water  then  runs  over  into  the  launder  at  m  and  descends  into  the  gang- 
way ditch  at  g  to  flow  to  the  sump;  whence  it  is  pumped  to  the  surface, 
where,  being  acid,  it  is  not  reused  unless  fresh  water  is  scarce.  At  the 
Dorrance  mine  where  the  rooms  were  being  filled  on  the  advance  by 
extending  the  flushing  pipe  from  one  panel  of  ten  rooms  to  the  next,  it 
was  the  practice  to  give  each  panel  another  dose  of  slush,  while  with- 
drawing the  pipe,  in  order  to  close  up  the  many  spaces  between  the  settled 
slush  and  the  roof  that  had  developed  since  the  advance.  For  nearly 
flat  seams,  dams  are  built  in  the  openings  all  around  a  panel  of  rooms, 
and  the  end  of  the  flushing  pipe  shifted  around  inside  the  panel,  close  to 
the  roof,  so  as  to  fill  all  portions  equally.  More  or  less  methane  is  given 
off  if  the  slush  is  exposed  to  air  currents,  but  these  are  feebler,  the  smaller 
the  spaces  left  between  slush  and  roof.  As  another  safeguard  against 
gas,  the  filled  panels  are  connected  with  the  return  airways  of  the 
active  mine. 

In  the  considerable  areas  where  a  subsidence  of  the  surface  is  imma- 
terial, the  anthracite  seams  are  best  worked  to  the  boundary,  by  that 
pillar  system  of  the  last  chapter  most  appropriate  to  the  given  conditions; 
and  the  pillars  then  recovered  on  the  retreat,  allowing  the  roof  to  fall. 
Under  the  river  flats  where  roof-falls  might  cause  a  crack  up  to  the  surface 
and  flood  the  workings,  one  company 's  mines  are  laid  out  with  permanent 
pillars  of  a  size  just  sufficient  to  sustain  the  roof  indefinitely,  which  means 
16-ft.  pillars  and  24-ft.  rooms  for  depths  of  less  than  400  ft. 

Flushing  as  a  preliminary  to  pillar-drawing  is  beneficial  in  the  anthra- 
cite region  under  two  conditions.  First,  where  the  workings  are  overlaid 
by  virgin  parallel  seams,  and  second,  where  they  are  overlaid  by  townsites. 
Former  market  conditions  made  the  thin  seams  unpayable,  so  that  the 
proper  method  of  exhausting  overlying  coal  seams  from  the  top  down- 
ward was  not  applied.  Now  the  pillars  can  only  be  recovered  from  the 
lower  seams,  without  wrecking  those  above,  by  a  preliminary  filling  of 
the  adjoining  rooms.  Formerly,  it  was  not  thought  that  the  pillars  left 
under  townsites  would  ever  be  worth  recovering,  but  higher  coal  prices 
have  made  them  valuable  and  filling  must  precede  their  recovery. 

The  aforementioned  Scranton  commission  recommends  that  as  slush 


282  MINING    WITHOUT    TIMBER 

alone  has  insufficient  crushing  resistance  for  thick  covers,  sand  should 
be  used  for  filling  under  Scranton  at  depths  beyond  500  ft.  Also  that 
filling  should  begin  in  the  lowest  seam  of  the  series  and  continue  upward 
until  all  are  filled,  care  being  taken  to  have  the  flushed  areas  over  one 
another.  After  all  the  openings  in  all  the  seams  have  been  filled,  the 
pillars  in  the  top  seam  may  be  removed  and  replaced  at  once  by  filling. 
The  next  seam  below  may  not  be  attacked  and  handled  in  like  manner 
until  the  pillars  above,  within  a  large  panel,  are  removed  and  the  over- 
burden has  come  to  rest  on  the  new  filling.  In  this  manner  several 
parallel  seams  could  be  robbed  of  pillars  simultaneously,  by  panels 
retreating  in  vertical  echelon,  the  robbing  in  the  highest  seam  being 
farthest  from,  and  that  in  the  lowest  seam  nearest  to  the  boundary. 

In  some  mines  with  irregular  layouts  and  small  pillars,  the  formation 
had  moved  considerable  before  flushing  was  inaugurated.  Thus  in  the 
22-ft.  Red  Ash  vein  of  the  Dodson  mine  at  Plymouth,  the  overlying 
formation  moved  so  freely  that  gangways  could  only  be  kept  open  by 
using  heavy  timbers  and  brushing  the  floor.  While  in  the  Black  Diamond 
mine  at  Luzerne,  the  walls  of  the  6-ft.  Cooper  seam  were  distorted  with 
frequent  roof -falls,  and  in  the  8-ft.  Bennett  seam 'the  roof  had  bent 
enough  to  badly  squeeze  many  of  the  pillars. 

The  seams  of  the  latter  mine,  which  were  excavated  on  the  system  of 
Fig.  130,  dip  about  10  deg.  and  the  pillars  of  the  flushed  portion  are  now 
being  robbed  and  replaced  by  slush.  Where  pillars  are  20  ft.  wide,  or 
more,  an  8-ft.  heading  is  driven  on  one  side  of  the  pillar  on  the  rise,  often 
leaving  a  thin  shell  of  coal  next  to  the  filling.  Then,  when  the  airway 
above  is  reached,  the  balance  of  the  pillar  is  drawn  on  the  retreat.  The 
advance  heading  must  be  well  propped,  but  the  timber  is  mostly  recov- 
ered on  the  retreat  and,  owing  to  the  moving  formation,  the  pillar  coal 
is  so  squeezed  that  but  little  blasting  is  necessary.  Too  much  roof  pres- 
sure sometimes  so  crushes  the  coal  that  it  falls  to  powder  when  extracted. 

The  Dorrance  mine  is  under  a  suburb  of  Wilkesbarre  and  the  policy 
of  the  owner,  the  Lehigh  Coal  Company,  is  to  refrain  from  taking  all  the 
pillar  coal,  when  robbing  flushed  areas  under  cities,  because  an  unsup- 
ported cover  will  settle  down  at  least  10  per  cent,  of  the  coal's  thickness; 
and  with  flat  seams,  where  filling  close  to  the  roof  is  impractical,  the  sub- 
sidence may  be  20  per  cent.  In  fact,  the  surface  in  some  cases  has  sub- 
sided less  from  robbing  pillars  in  open  than  in  filled  seams;  for  in  the 
former  case  local  breaks  of  roof  may  fill  up  the  rooms  with  boulders  and 
support  the  cover,  while  robbing  pillars  completely  in  the  latter  case 
starts  the  whole  cover  to  subsiding  as  in  the  longwall  system. 

The  flushed  workings  observed  in  the  Dorrance  mine  were  on  a  dip 
of  18  deg.  and  on  a  layout  like  Fig.  130  with  rooms  20  ft.  and  pillars  40  ft. 
wide.  A  heading  was  first  driven  up  in  the  pillar,  to  slab  off  18  ft.  of 
coal  alongside  the  filling,  and  on  the  retreat  from  the  room 's  upper  end 


FLUSHING   SYSTEMS   FOR   FILLING    SEAMS  283 

a  24-ft.  cross-cut  was  put  through  the  pillar,  halfway  between  the  original 
12-ft.  break-throughs,  100  ft.  apart.  Thus  after  flushing  the  new  pillar 
openings,  the  roof  was  left  supported  by  a  line  of  coal  pillars  22  feet 
\\ide  by  32  ft.  along  the  dip. 

Under  Mahonoy  City,  the  22-ft.  Mammoth  seam,  dipping  55  to 
60  deg.,  is  being  worked  in  two  slices  by  a  system  like  that  of  Fig.  131. 
The  lower  slice  of  15  ft.  is  taken  out  in  the  room  on  the  advance  and  the 
upper  7-ft.  slice  allowed  to  fall  into  the  chute,  by  pulling  the  props,  on 
the  retreat.  After  flushing,  the  pillar  is  taken  out,  likewise,  in  two 
slices,  by  driving  a  heading  through  its  center,  leaving  only  a  thin  shell 
of  coal  on  each  side  to  keep  out  the  room-filling.  The  entry  pillars  are 
drawn  on  the  retreat,  and  all  the  new  openings  are  flushed.  In  spite  of 
this  extraction  of  practically  all  the  pillars,  the  surface  here  is  stable,  for 
with  seams  of  steep  dip,  the  subsidence  upon  the  filling  is  not  so  serious 
as  it  is  in  the  case  of  the  flatter  seams  under  Wilkesbarre. 

As  already  mentioned,  the  22-ft.  seam  in  the  Dodson  mine  is  also 
worked  in  two  slices  but  with  the  thin  slice  below.  The  pillars  here  are 
26  ft.  and  the  room  is  24  ft.  wide.  The  lower  slice  of  both  room  and  pillar 
is  mined  on  the  advance  and  the  upper  slice  is  recovered  on  the  retreat 
as  described  in  the  last  paragraph,  except  that  the  layout  follows  Fig.  130 
to  suit  the  12-deg.  dip. 

EXAMPLE  60. — ROBINSON  GOLD  MINE,  RAND  DISTRICT.     TRANSVAAL 
Parallel  Sloping  Beds  Filled  with  Mill  Tailing 

In  spite  of  the  extensive  areas  excavated  since  1885  in  the  conglom- 
erate of  the  Rand,  but  little  filling  has  yet  been  done.  At  a  few  rich 
outcrop  mines,  it  is  true,  the  rooms  were  packed  with  rock  to  enable  the 
pillars  to  be  recovered.  But  packing  is  too  costly  a  method  for  most 
of  the  area.  As  the  mines  reach  depths  exceeding  4000  ft.,  the  former 
sized  pillars  are  proving  too  small,  and  several  unexpected  collapses 
have  occurred.  Recently  the  flushing  system  has  been  tried  with  suc- 
cess at  the  Robinson  mine,  to  permit  of  the  removal  of  some  rich  pillars 
just  under  the  stamp  mill,  in  the  following  manner. 

The  tailing  is  washed  from  the  dump  by  a  1-in.  water  pipe  into  a 
launder,  6  in.  sq.,  which  runs  to  the  top  of  a  winze.  Here  the  pulp  enters 
a  similar  launder  which  descends  along  the  40-  to  50-deg.  dip  of  the  seam 
floor  to  the  ninth  level  of  the  mine.  The  stope  to  be  filled  has  been 
dammed  at  the  lower  end  by  a  dry  wall  W,  see  Fig.  145,  strengthened 
by  poles  P,  and  similar  partition  walls  are  built  at  right  angles  to  cut  it 
up  into  longitudinal  panels.  The  fine  waste  B  is  piled  above  W,  and 
covered  with  old  matting  M  from  the  cyanide  tanks.  When  flushing 
begins,  the  sand  settles  quickly,  the  water  filters  through  the  matting 
and  dams,  whence  it  runs  to  the  sump  to  be  pumped  to  the  surface. 


284 


MINING    WITHOUT    TIMBER 


This  water  is  used  again  after  a  little  lime  has  been  added  to  neu- 
tralize its  acidity  and  render  any  entrained  colloids  harmless  to  hinder  a 
quick  settling.  To  save  water,  the  launders  are  kept  on  a  minimum 
gradient  of  10  deg.  The  water  used  is  6  to  10  per  cent,  of  the  tailing  by 
weight.  The  cost  of  filling  is  given  at  2.1  d  per  ton,  but  as  only  100  tons 
of  tailing  are  sent  underground  daily,  this  probably  does  not  include 
wear  of  the  launders.  The  filling  sets  hard  in  2  or  3  days.  When  a  stope 
is  completely  filled,  it  only  settles  10  per  cent,  of  its  height  when  crushed 
by  the  formation  after  the  pillars  have  been  removed. 

The  residual  cyanide  of  the  tailing  leaving  the  mill  has  been 
destroyed  by  exposure  on  the  old  dumps,  so  that  no  poisonous  results 
have  so  far  ensued  from  using  tailing  as  mine  filling.  In  order  to  utilize 
fresh  tailing,  the  cyanide  must  first  be  rendered  innocuous.  This  is  not 


Cross  Sec. 


FIG.  145. — Dam  for  holding  slush,  Transvaal. 

urgent  at  present,  because  the  old  tailing  dumps  are  immense.  Flushing 
the  leading  vats  direct  into  the  mine,  however,  would  save  the  expense 
of  conveying  the  tailing  to  the  top  of  the  very  high  dumps  and  of  redig- 
ging  it  before  flushing.  Hence  some  mines  are  now  getting  ready  for 
direct  flushing. 

The  flushing  system  is  now  being  freely  used  in  the  Rand  to  fill  stopes 
not  under  buildings,  in  order  to  prevent  the  damage  to  the  workings  and 
the  shaft  pillars  which  is  liable  to  ensue  from  pillar-drawing,  especially 
as  the  mines  get  deeper.  The  rock  tailing  available  for  filling  is  much 
more  resistant  to  crushing  than  the  anthracite  refuse  of  Example  59, 
and  is  strong  enough  for  a  filling  at  any  workable  depth.  On  the  central 
Rand,  there  are  two  contiguous  parallel  beds,  the  Main  Reef  below,  and 
the  Main  Reef  Reader  above,  separated  by  a  thin  rock  parting.  As  the 
Main  Reef  is  the  leaner,  it  has  hitherto  been  neglected  in  many  mines,  but 
it  is  expected  that  the  flushing  system  will  now  greatly  facilitate  its 
extraction  under  the  worked-out  stopes  of  the  Main  Reef  Leader. 


CHAPTER  XXII 
COMPARISON  OF  VARIOUS  MINING  SYSTEMS 

The  grouping  of  the  practical  Examples  of  this  book  has  been  based 
on  the  system  described  in  each  case  in  the  general  title  of  the  chapter, 
while  each  example  has  also  an  individual  title  suggesting  its  own  dis- 
tinctive features.  Mining  being  an  applied  rather  than  a  pure  science,  the 
problem  of  choosing  an  appropriate  system  for  any  mine  is  a  commercial 
question.  Nevertheless,  the  problem  should  be  viewed  in  the  broad 
instead  of  the  narrow  sense.  Greater  temporary  profits  may  mean  lesser 
ultimate  profits  for  the  whole  property  owing  to  loss  of  ore.  A  given 
system  may  be  cheap  to  operate  but  dangerous,  and  the  cost  of  resulting 
accidents  may  greatly  over-balance  any  temporary  gains. 

Besides  the  question  of  ore  recovery  and  safety  for  life  and  property, 
the  choice  of  mining  systems  involves  such  mechanical  engineering 
problems  as  those  of  hoisting,  haulage,  pumping,  ventilation,  and  lighting, 
and  such  miners'  problems  as  those  of  breaking  ground  and  controlling 
excavations  under  given  conditions.  Other  things  being  equal,  that 
system  is  preferable  which  minimizes  the  dead  work  for  shafts,  cross-cuts, 
and  raises;  which  breaks  the  ore  with  the  least  drilling  and  explosives, 
and  which  keeps  the  mine  open  with  the  least  artificial  support.  A 
system  must  adapt  itself  to  the  district's  labor,  whether  skilled  or  un- 
skilled, cheap  or  dear,  scarce  or  plentiful,  and  must  fit  the  mineral  market 
conditions,  whether  steady  or  fluctuating.  If  shut-downs  are  likely  to  be 
periodic  from  strikes  or  glutted  markets,  the  mining  system  chosen 
must  allow  the  mine  to  lie  idle  with  a  minimum  damage  to  its  workings. 

As  the  advantages  and  disadvantages  of  each  Example  have  already 
been  given  with  its  description,  only  a  general  comparison  need  be  made 
here  between  the  different  Examples. 

The  drag-line  excavators  of  Cuba  described  in  Chapter  VI  are  only 
superior  to  steam  shovels  for  deposits  having  a  rough  bottom  on  which 
the  cost  of  track-laying  for  the  latter  system  would  not  be  justified  by 
the  depth  of  the  ore  to  be  shovelled  from  a  given  trackage.  The  shal- 
lower deposits  of  the  Mesabi  Range  and  of  Ely,  Nev.,  are  well  adapted  to 
steam-shovel  work,  as  neither  the  top  nor  the  bottom  of  the  ore  lenses 
are  sufficiently  irregular  to  offer  serious  difficulties  in  reaching  all  ore  or  in 
laying  tracks,  and  dumping  ground  for  stripping  is  nearby.  At  Bingham, 
Utah,  however,  natural  conditions  do  not  favor  steam  shoveling  because 
the  top  of  the  ore  deposit  is  nearly  parallel  to  the  surface  of  a  precipitous 

285 


286  MINING    WITHOUT   TIMBER 

mountain  and  not  only  is  it  difficult  to  shovel  stripping  and  ore  separately, 
but  much  expensive  grading  is  required  for  track  connections  to  the 
numerous  shovel  benches.  It  is  not  likely  that  steam  shovels  will 
again  be  installed  in  such  rough  topography  as  that  of  Bingham,  for  sur- 
face-milling for  thin,  and  underground-caving  for  thick,  capping  is  far 
preferable.  For  properly  laying  out  open  cuts,  preliminary  prospecting 
is  especially  important;  therefore  its  various  details  have  been  carefully 
described  in  Chapter  VI.  The  opencut  steam-shovel  work  of  Example  5 
avoids  all  the  dangers  and  most  of  the  expense  of  underground  coal 
mining  and  is  eminently  suitable  for  all  flat  seams  with  thin  covers;  it 
has  been  used  for  the  excavation  of  the  Clinton  iron  ore  bed  of  New 
York. 

The  open  quarry  of  Puertocitos  of  Chapter  VII  is  well  suited  to  a 
sidehill  location,  but  for  other  sites  the  milling  system  avoids  the  shovel- 
ing of  the  broken  rock  and  should  be  adopted  where  the  ore  body  is 
large  enough  to  warrant  the  cost  of  the  preliminary  development  neces- 
sary. The  hour-glass  chutes  of  Example  9  are  especially  suitable  to 
work  on  a  large  scale  and  in  winter,  for  the  smaller  boulders  can  be  blasted 
at  leisure  in  the  warm  chute  with  no  danger  to  -the  pit  men  from  flying 
pieces. 

In  Chapter  VIII,  the  difference  in  the  systems  of  Examples  10  and  11 
are  slight  and  are  due  to  the  fact  that  flat-holes  are  better  adapted  to  the 
Joplin  formation  than  the  usual  down-holes  of  quarrying.  It  is  evident 
that  deposits  are  only  adapted  to  underhand  stoping  when  they  have 
considerable  vertical  height.  Flat  deposits  must  have  a  strong  roof  and 
sub-vertical  veins  must  be  of  ore  sufficiently  strong  to  form  a  self-sustain- 
ing back  over  a  large  stope,  unless  a  saddle-back  support  of  timber  can 
be  rigged  up  as  suggested  for  Example  13.  Unless  spots  of  lean  ore 
occur  so  as  to  be  utilizable  for  pillars,  the  ordinary  underhand  systems 
may  involve  considerable  loss  of  good  ore  in  pillars,  excepting  where 
auxiliary  back-caving  can  be  used  as  described  in  Example  13.  The 
Mitchell  system  is  an  attempt  to  adopt  underhand  stoping  to  soft  ground, 
by  combining  it  with  square  setting.  This  system  has  promise  of  con- 
siderable usefulness  in  suitable  formations. 

Among  the  advantages  of  underhand  stoping  are  the  cheap  drilling 
and  breaking  that  go  with  underhand  benches  and  the  fact  that  no  broken 
ore  is  locked  up  in  the  stopes  as  in  all  shrinkage  systems,  or  lost  or  con- 
taminated by  mixing  with  waste  as  in  the  filling  and  caving  systems. 
Timbering  and  developing  work  is  at  a  minimum  in  underground  quarry- 
ing but,  as  an  offset,  all  broken  ore  has  to  be  loaded  by  shoveling,  which 
disadvantage  may  be  overcome  by  the  milling  system  of  Example  13. 
In  stoping  veins  only  one  stope  can  be  worked  at  a  time  on  one  level  on 
each  side  of  the  shaft,  but  several  levels  can  be  stoped  at  once  by  keeping 
the  top  ones  farthest  advanced  and  leaving  a  longitudinal  pillar,  or  at 


COMPARISON    OF   VARIOUS    MINING    SYSTEMS  287 

least  a  shelf,  of  ore  under  each  level  to  support  the  track  over  the  stopes 
beneath.  The  underhand  systems  give  no  space  for  stowing  waste  in 
the  stope  where  it  is  broken  and  are  therefore  best  adapted  to  clean  ore. 

The  first  four  examples  of  the  simpler  overhand  stoping  systems  of 
Chapter  IX  are  suited  to  strong  ores  and  wall  rocks.  In  both  the 
Wolverine  and  the  Homestake  systems,  all  the  broken  ore  has  to  be  loaded 
by  shoveling,  but  this  has  been  preferred  to  the  greater  initial  expense  of 
building  chutes  for  whose  smooth  working  the  Wolverine  footwall's  dip 
is  scarcely  sufficient.  The  use  of  triangular  ore  pillars  or  "rills"  above 
the  drifts,  is  an  ingenious  device  to  save  timber.  They  can  be  as  easily 
recovered  when  the  level  is  abandoned  as  pillars  in  other  places.  By 
leaving  enough  broken  ore  or  " shrinkage"  close  to  the  face,  money  is 
locked  up  it  is  true,  but  the  expense  of  timber  scaffolds  is  saved  for  the 
miners  in  steep  veins.  The  broken  ore  also  acts  as  a  temporary  support 
to  prevent  caving  in  case  the  stope  walls  are  weak  or  the  roof  pressure  is 
great.  By  using  overhead  chutes  for  filling  the  cars,  a  line  of  stopes  can 
be  worked  simultaneously  along  one  level,  but  this  system  gives  no  space 
for  storing  waste  in  the  stope  where  broken.  The  stoping  system  of 
Example  18  has  panel-cores  inside  of  the  square  set  frames  like  Example 
12,  but  its  method  of  attack  is  quite  different  and  the  system  of  the 
former  is  less  well  adapted  to  soft  ground. 

Where  the  walls  are  so  weak  as  to  cave  on  emptying  the  stope  or 
where  no  ore  can  be  spared  for  pillars  and  where  much  broken  vein  matter 
is  worthless,  the  systems  of  Chapters  X  and  XI,  which  use  permanent 
waste  filling  instead  of  the  shrinkage  of  Chapter  IX,  are  applicable.  In 
the  American  examples  of  Chapter  X,  the  dry-walled  drifts,  the  descend- 
ing hangwall,  the  rill  chutes  and  the  auxiliary  milling  and  square  setting 
are  noteworthy  for  Examples  19,  20,  21  and  22  respectively. 

For  the  soft  ore  bodies  of  Bisbee  with  their  irregular  shape  and 
their  uncertain  boundaries,  square-setting  in  panels  seems  to  be  especially 
suitable  for  the  reasons  given  in  the  description  of  Example  23.  The 
soft  walls  and  the  fact  that  a  large  portion  of  the  broken  material  has 
to  be  left  in  the  stope  as  waste  precludes  the  use  of  any  caving  system 
except  perhaps  one  which,  like  Example  43,  is  well  timbered  and  allows 
of  stowage  near  the  face. 

Considering  the  foreign  systems  of  Chapter  XI,  a  similarity  can  be 
seen  between  the  Australian  Example  25  and  the  American  Example  21, 
and  between  the  Mexican  Example  23,  the  Australian  Example  26,  and 
the  American  Example  22.  In  Example  27,  the  fragmentary  nature  of 
the  vein  filling  has  necessitated  a  system  where  no  opening  need  be 
made  wider  than  a  cross-cut. 

Not  only  does  the  adoption  of  rills  save  the  cost  of  timbering  the 
back  of  the  drift  where  the  ore  is/strong,  but  it  saves  much  labor  as  the 
broken  ore  slides  down  to  the  discharge  chute  and  the  filling  from  the 


288  MINING    WITHOUT   TIMBER 

level  above  slides  from  the  delivery  winze  to  its  place  with  little  or  no 
handling.  For  the  filling  systems,  the  rills  should  have  a  dip  of  37  deg.  to 
permit  the  waste  to  slide  continually  and  freely,  while  for  the  shrinkage 
systems,  where  they  are  used  as  slides  but  once,  the  rills  need  dip  only  23 
deg.  Filling  with  "flat-back"  chutes,  where  the  stope  is  carried  horizon- 
tally and  the  drift  roofed  with  timber  for  its  whole  length,  has  -the 
advantage  over  rill  chutes  in  heavy  ground  of  a  filling  whose  top  surface 
is  horizontal.  This  feature  offers  a  better  base  for  timber  cribs  or  sets 
to  support  a  weak  back. 

In  comparing  shrinkage  with  waste-filling  by  rills,  the  former  needs 
only  about  one-fourth  the  number  of  winzes  from  the  level  above  as 
the  latter,  and  none  of  the  expensive  ore-passes  of  filling  are  required 
for  shrinkage  whose  ore  is  always  drawn  off  at  the  bottom  of  the  stope. 
Also  air  drills  do  not  have  to  be  continually  shifted  in  shrinkage  to 
allow  the  filling  to  be  placed.  Nor  is  any  ore  lost  by  shifting  down  into 
waste.  Nevertheless,  shrinkage  ore  is  often  more  or  less  contaminated 
by  scaling  off  of  the  walls  when  emptying  the  stope. 

Ore  bodies  with  walls  hard  enough  not  to  cave  when  the  shrinkage  is 
withdrawn,  and  containing  ore  clean  enough  to  be  hoisted  without  sort- 
ing, are  best  suited  to  the  systems  of  Chapter  XII  which  combine  shrink- 
age and  waste  filling.  Example  28  uses,  where  the  ore  is  weak  or  irreg- 
ular, auxiliary  cribbing  and  waste  filling.  Both  Examples  29  and  30  use 
some  back-caving  to  cheapen  the  breaking  of  the  ore.  The  latter  finds 
some  underhand  stoping  advantageous  in  breaking  down  weak  and 
dangerous  stope-backs.  Example  30  is  the  only  one  which  is  now  ex- 
tracting its  pillars  but  there  is  no  reason  why  Examples  28,  29  and  30 
should  not  also  remove  any  pillars  of  ore  by  slicing,  after  the  adjoining 
rooms  have  been  filled  with  waste. 

The  systems  of  Chapter  XIII,  where  stoping  and  shrinkage  in  the 
rooms  is  combined  with  caving  of  the  pillars,  may  be  considered  hybrids 
between  shrinkage  and  the  true  caving  systems  of  Chapters  XIV 
to  XVII.  By  making  the  rooms  of  the  systems  of  Chapter  XIII  quite 
narrow,  we  could  have  a,  block-caving  system  with  the  rooms  serving 
as  mere  cut-off  stopes.  The  Miami  system  resembles  that  of  Example 
30  except  that  the  latter  waits  to  fill  a  room  before  slicing  an  adjoining 
pillar,  while  the  former  withdraws  pillar  and  shrinkage  simultaneously 
and  avoids  any  filling  whatever.  Mr.  Lawton's  idea  of  keeping  the 
weak  ore  of  a  60-ft.  room  supported  by  a  sharply  arched  back  whose 
haunches  rest  on  broken  ore  blown  under  them  by  specially  placed 
holes,  is  a  great  saving  over  the  cribs  under  the  back  of  Example  26. 

The  Boston-Con,  system  resembles  the  Miami  in  the  stoping  rooms, 
but  its  failure  to  similarly  support  the  haunches  of  the  weak  back  with 
broken  ore  caused  its  working  to  be  dangerous  for  the  miners.  The 
Duluth  system  resembles  the  Boston-Con,  but,  the  ore  being  strong- 


COMPARISON    OF   VARIOUS   MINING    SYSTEMS  289 

enough  to  stand  over  the  rooms,  there  is  not  the  same  danger  in  its 
operation  from  falls  of  the  back.  Conversely,  to  get  the  pillars  to  cave 
sufficiently  for  passing  a  chute,  they  have  to  be  kept  thin.  The  pillars 
when  block-caved,  are  evidently  less  under  control  than  when  sliced 
as  at  Miami,  and  the  saving  of  the  timber  and  breaking  costs  of  slicing, 
may  be  offset  by  greater  loss  of  ore,  more  pollution  by  the  descending 
capping  and  more  hung-up  chutes. 

As  already  noted,  back-caving  is  an  auxiliary  in  the  overhand  stoping 
of  Examples  29  and  30,  but  the  chute-caving  of  Chapter  XIV  is  an 
attempt  to  break  all  the  ore  by  back-caving  except  that  taken  out  by 
necessary  development  openings.  In  the  Hartford  system  the  only 
development  needed  in  stoping  is  that  of  the  spiral  raises  around  the 
caving  cores  of  the  back.  This  system,  however,  can  only  be  successful 
with  well  defined  and  strong  walls  and  a  regular  deposit  of  clean  ore. 
By  the  use  of  sub-levels  in  Examples  36  and  37,  chute-caving  becomes 
safer,  more  easily  controlled  and  better  adapted  to  irregular  deposits 
with  weak  walls,  though  even  with  them  there  is  no  place  to  store 
waste  in  the  stops.  The  subsequent  uncovering  by  the  steam-shovel 
work  of  a  caved  portion  of  the  Utah-Copper  mine  showed  a  heavy  loss 
of  ore  from  mixtures  with  capping.  Much  of  this  loss,  however,  was 
due  to  unsystematic  work  and  is  not  inherent  in  chute-caving  itself. 

The  block-caving  of  Chapter  XV  allows  the  cheapest  mining  of  any 
underground  system,  yet  it  is  liable  to  the  heaviest  loss  of  ore  with 
unsuitable  deposits.  The  bigger  the  block,  the  less  liable  is  the  detaching 
drop  to  crush  it  completely.  Success  at  the  Pewabic  is  due  to  the  fact 
that  the  ore  is  sandy  and  friable.  At  other  Lake  Superior  iron  mines  with 
hard  and  tough  ore,  block-caving  proved  a  failure  because  only  the 
lower  part  of  the  mass  was  crushed  and  the  solid  core  would  not  descend 
to  be  loaded.  Although  the  Pewabic  spstem  has  no  chutes,  which  are 
expensive  to  construct  and  likely  to  clog,  it  has,  as  an  offset,  to  cross-cut 
the  broken  ore  of  the  block  and  load  it  all  by  hand-shoveling. 

The  Mowry  ore  being  friable  and  located  in  a  small  vertical  pipe 
offers  ideal  conditions  for  block-caving,  and  the  plan  of  placing  a  square- 
settled  floor  at  the  base  of  each  block,  saves  the  hand-loading  expense 
of  the  Pewabic.  In  Examples  40  and  41  the  blocks  are  much  smaller 
than  at  the  Pewabic.  Consequently  they  can  be  crushed  sufficiently  by 
one  drop  to  be  drawn  through  chutes.  Conversely,  as  much  breaking- 
may  be  required  as  with  chute-caving  at  the  Utah-Copper  where  it  is 
reckoned  that  30  per  cent,  of  the  ore  is  extracted  by  development  work. 

On  drawing  caved  ore  from  chutes,  it  is  essential  to  provide  means 
for  reaching  the  top  of  the  chutes  in  order  to  break  up  any  boulders  that 
cause  stoppage.  This  has  been  arranged  for  in  the  various  examples 
cited  so  to  save  delay  in  tramming  and  an  expensive  interference  with 
the  mine's  routine.  At  the  Inspiration  mine,  the  chutes  are  expanded 

19 


290  MINING    WITHOUT   TIMBER 

into  drawing-off  slopes  and  have  been  made  the  central  feature  of  the 
system.  These  chute-stopes  are  ingeniously  arranged  to  aid  the  self- 
crushing  of  the  descending  block  and  thus  enable  larger  blocks  and 
tougher  ore  to  be  handled.  The  proposed  use  of  a  mat  of  ore  is  also  an 
innovation  and  should  greatly  decrease  the  loss  of  ore  which  occurs 
when  the  main  block  breaks  partly  into  boulders  which  may  hang  up  in 
descending  and  permit  the  sliding  beneath  of  quantities  of  broken 
capping.  Timber  mats  are  worthless  for  block-caving  because  they  are 
ground  up  by  settling  and  there  is  no  chance  to  repair  them  as  in  the 
case  of  slicing.  A  mat  of  ore,  on  the  contrary,  breaks  into  boulders  be- 
fore the  main  block  beneath  does  and  prevents  any  waste  descending 
into  the  interstices  of  the  latter  during  its  slow  disintegration.  The 
higher  the  caved  block,  the  less  proportional  loss  from  pollution  of  the 
top  layer  with  the  descending  capping,  so  that  at  the  Inspiration,  with 
a  block  over  200  ft.  high,  it  is  hoped  to  reduce  the  loss  of  ore  to  under 
15  per  cent. 

In  the  slicing  systems  of  Chapter  XVI  and  XVII,  the  ore  is  broken 
entirely  by  breast  stoping  instead  of  by  the  underhand  or  overhand 
stoping  of  previous  chapters.  The  "slicing  under  mats"  of  Chapter  XVI 
takes  almost  as  much  timber  as  square-setting,  but  an  inferior  quality 
may  be  used.  Slicing  has  also  the  disadvantages  of  continually  breaking 
the  ore  from  the  bottom  under  the  heavy  pressure  of  the  superincumbent 
filling,  and  of  poor  ventilation.  But  it  allows  the  use  of  a  larger  propor- 
tion of  muckers  to  miners  and  a  larger  output  per  man  than  square- 
setting,  and  is  not  liable  to  the  collapses  to  which  the  latter  is  subject 
in  heavy  ground  unless  closely  filled.  Considerable  ore  may  be  lost 
by  a  little  carelessness  in  square-setting  by  its  sifting  into  the  filling 
beneath,  but  in  slicing  any  such  siftings  fall  onto  solid  ore  and  are  re- 
covered later. 

The  "slicing  under  ore  in  rooms"  of  Chapter  XVIII  resembles  the 
"slicing  under  mats  in  panels"  of  Chapter  XVII,  but  there  is  a  large 
saving  in  the  consumption  of  powder  and  of  timber  at  the  expense  of  a 
slightly  greater  loss  of  ore  which  should,  however,  never  exceed  a  total 
of  5  per  cent.  The  use  of  rooms,  one  set  wide,  rather  than  broad  panels 
allows  the  use  of  rough  unframed  timber,  and  the  recovery  of  a  half-set 
width  of  ore  on  each  side  and  outside  of  the  room-set  without  any 
timber  whatever.  "Slicing  under  ore"  takes  less  skill  and  is  quicker 
than  "slicing  under  mats."  Moreover,  the  timber  floor  mats  of  the 
former  system  do  not  have  to  be  nearly  as  heavy  or  as  well  interlaced 
as  those  of  the  latter.  As  can  be  seen  in  the  given  examples,  either 
system  is  flexible  and  applicable  to  various  classes  of  deposits,  but  in 
ore  bodies  of  any  size  or  irregularity,  "slicing  under  ore"  is  far  preferable 
excepting  where  the  ore  is  too  valuable  to  permit  of  any  loss  at  all,  or 
where  it  contains  much  waste,  for  only  "slicing  under  mats"  permits 


COMPARISON   OF   VARIOUS   MINING   SYSTEMS  291 

stowage  of  waste  in  the  stope  where  broken.  At  the  Kimberley  mine, 
the  ore  is  sliced  and  back-caved  in  galleries  without  the  use  of  timber. 
This  system  is  applicable  to  any  large  deposit  where  the  ore  is  homo- 
geneous and  of  sufficient  firmness. 

The  nature  of  both  capping  and  ore  affect  the  choice  of  a  caving 
system.  A  capping  which  breaks  large  and  chunky  is  favorable  to 
block  or  pillar-caving,  for  fine  capping  sifts  down  into  the  large  spaces 
in  the  disintegrating  ore  blocks  and  pollutes  the  ore.  On  the  contrary, 
friable  capping  favors  chute-caving  and  slicing  because  the  roof  settles 
more  evenly  in  these  systems.  In  slicing,  any  chunks  above  would 
tend  to  break  through  the  timber  mat  every  time  one  floor  was  dropped 
10  ft.  to  the  next  below. 

In  dropping  a  block  of  ore  for  caving,  it  is  crushed  both  by  the 
superincumbent  weight  and  by  attrition  on  the  side  walls.  In  "slicing 
under  mats,"  all  the  ore  is  broken  by  explosives,  but  in  the  other  systems, 
greater  or  less  proportion  is  crushed  by  dropping.  It  is  evident  that 
the  larger  blocks  and  the  shorter  and  fewer  drops,  the  less  crushing 
action  is  exerted  on  the  ore.  The  order  of  crushing  efficiency  for  the 
different  systems  would  generally  be  "  slicing  under  ore,"  chute-caving, 
pillar-caving  and  block-caving.  The  last  named  system  would  re- 
quire the  most  friable  ore  of  all  in  order  to  be  successful.  If  the  capping 
looks  different  from  the  ore,  it  is  separated  more  easily  when  drawing 
down  a  chute.  Fortunately,  this  is  usually  the  case  in  the  examples 
quoted  for  if  not  of  a  different  color,  the  capping  is  more  silicified  and 
breaks  in  larger  pieces  than  the  ore  and  differently  shaped. 

In  adopting  a  caving  system  for  a  mine,  after  due  consideration  has 
been  given  the  size,  shape,  grade,  location,  etc.,  of  the  ore  body,  a  person 
should  be  able  to  narrow  the  final  choice  to  two  possible  systems.  One 
of  these  systems  will  mean  a  higher  mining  cost  along  with  a  higher 
percentage  of  ore  saved  than  the  other.  If  the  net  commercial  result 
is  practically  the  same  in  both  cases,  it  is  usually  better  to  choose  the 
higher  cost  system  not  only  for  reasons  of  conservation  of  resources, 
but  because  any  future  increase  in  the  market  value  of  the  ore  will  be 
of  most  advantage  to  the  system  making  the  best  saving.  Of  the  caving 
systems,  block-caving  should  usually  cost  the  least.  Then  follow  pillar- 
caving,  chute-caving,  slicing-under-ore  and  slicing-under-mats  in  the 
order  named.  The  percentage  of  saving  is  in  inverse  order,  but  indi- 
vidual conditions  may  change  the  order  of  cost  and  saving,  respectively, 
to  a  consider  :b'e  degree. 

The  continuous-face  longwall  system  of  Example  49  is  adapted  to 
flat  seams  with  little  gas.  The  oblique  angled  roads  which  are  the  dis- 
tinctive feature  of  the  Scotch  system  are  here  at  45  deg.,  but  have  other 
angles  (up  to  60  deg.)  elsewhere.  In  Grundy  County,  mines  have  been 
worked  by  the  Rectangular  system,  with  all  roads  at  right  angles,  but, 


292  MINING    WITHOUT    TIMBER 

although  the  latter  makes  it  simpler  to  maintain  a  regular  face,  it  re- 
quires longer  haulageways  and  is  therefore  less  in  favor.  Where  the 
seam  is  sloping  or  where  much  gas  may  make  it  advisable  to  specially 
regulate  or  shut  off  the  ventilation  of  a  dangerous  locality,  a  panel  sys- 
tem is  advantageous.  The  limit  of  dip  for  panel-longwall  is  about 
40  deg.,  for  seams  with  yet  steeper  dips  the  systems  in  vogue  follow  over- 
hand stoping,  either  "with  shrinkage"  or  "on  waste,"  as  described  in 
Chapters  IX  to  XI. 

Example  50  gives  the  advancing,  panel-longwall  method  for  stoping 
iron  beds.  Although  the  method  was  in  general  use  decades  ago,  it  is 
still  applicable  to  ore  beds  that  present  the  above-mentioned  conditions 
favorable  for  longwall.  It  is  especially  suited  to  dips  steep  enough  to 
allow  the  broken  ore  to  descend  by  gravity  down  the  footwall  into  the 
chutes.  Example  51  illustrates  how  the  panel  system  can  be  applied  to 
a  seam  with  an  irregular  floor  and  a  cracked  roof,  features  which  are  often 
stated  to  be  fatal  to  successful  longwall.  The  handling  of  the  coal  along 
the  face  by  a  buggy  on  an  endless  rope  allows  the  gateways  to  be  125  ft. 
apart  instead  of  the  42  ft.  of  Example  49  or  the  20  to  30  ft.  of  Example  50, 
and  thus  saves  a  proportional  expense  in  the  building  and  maintenance 
of  roadways.  . 

Unlike  the  three  previous  examples,  Examples  52  and  53  consume 
much  timber,  but  this  defect  is  not  inherent  in  the  systems  but  is  due  to 
the  fact  that  neither  from  the  partings  in  the  seams  nor  from  the  strata 
of  the  roof  was  sufficient  material  secured  for  stone  pack-walls  and  cogs. 
The  Vintondale  conveyors  are  certainly  labor  savers  for  thin  veins,  and 
where  their  liability  to  frequent  breakdown  is  obviated  the  present  dislike 
of  the  miners  for  the  system  can  easily  be  overcome.  The  "  jig  "  roads  at 
Westville  are  well  adapted  to  the  average  22  1/2-deg.  dip,  and  the  41-ft. 
width  of  room  is  nearly  the  same  as  for  the  similar  loading  in  Example  49. 
Westville  illustrates  that  for  deep  mines  with  weak  coal,  longwall  is  the 
only  practical  system  and  there  should  be  no  difficulty,  when  the  upper 
7-ft.  layer  has  been  extracted  and  the  roof  has  come  to  rest  on  its  floor, 
in  similarly  removing  the  lower  10-ft.  of  the  seam. 

The  various  figures  of  Example  54  illustrate  merely  the  simpler  outlines 
of  the  pillar  system.  To  suit  all  the  conditions  of  different  mineral  districts, 
an  endless  variety  of  layouts  has  been  devised  for  modern  mines.  The 
entries  vary  from  2  to  8  headings  abreast,  and  the  angles  at  which  the 
butt  and  cross  entries  meet,  and  at  which  the  latter  meet  the  main  entries, 
is  often  made  acute,  to  suit  the  dip,  instead  of  90  deg.  The  ratio  of  room- 
width  to  pillar-width  is  varied  to  suit  the  formation,  the  depth  of  cover, 
the  seam,  the  method  of  mining  and  the  desired  size  of  mineral.  The 
selection  of  mechanical  equipment  and  the  coupling  of  entries  and  shafts 
for  effective  and  safe  ventilation  and  for  rapid  movement  of  a  large  output 


COMPARISON    OF   VARIOUS    MINING    SYSTEMS  293 

of  mineral,  especially  at  the  shaft  bottom,  are  problems  requiring  special 
treatment  for  each  case. 

The  Nelms'  layouts  of  Examples  55  and  56  are  well  arranged  to  pro- 
duce a  large  product  from  a  small  area.  Such  intensive  mining,  when 
thus  systematically  carried  on,  conduces  not  only  to  safety  but  also  to 
the  minimum  cost  of  equipment  and  operation. 

Examples  57  and  58  illustrate  methods  far  in  advance  of  much  of  the  bi- 
tuminous mining  practise  of  America.  The  thickness,  fine  quality  and  ac- 
cessibility of  thePittsburg  seam  has  given  it  a  value  ranging  up  to  $3,000 
an  acre,  and  this  fact  has  tended  to  promote  scientific  working  and  a  high 
percentage  of  recovery.  In  Example  57,  illustrating  mining  for  coking 
coal,  the  first  method  of  recovering  pillars  without  the  use  of  curves  is 
suited  for  narrow  pillars,  while  the  two  last  curve-methods  are  suitable 
to  the  wider  pillars  necessary  under  heavy  covers.  The  three  methods 
described  in  Example  58  are  those  for  the  market-coal  mines  and  are 
designed  to  produce  a  maximum  of  lump  coal  with  the  use  of  machine 
cutters.  They  are  laid  out  for  a  large  output  and  with  careful  work  make 
possible  a  total  recovery  of  90  per  cent,  of  the  seam. 

The  flushing  system  of  Chapter  XXI  has  been  a  boon  to  coal  mining 
in  thickly  populated  districts.  It  is  really  nothing  but  the  old  European 
filling  system,  using  fine  surface  waste  for  filling  and  putting  it  under- 
ground without  any  of  the  manual  labor  that  makes  the  old  system  so 
expensive.  The  cleaning  up  of  the  great  culm  heaps  in  the  anthracite 
region,  by  washing  and  flushing,  has  rendered  usable  large  areas  of  valu- 
able surface  land  as  well  as  made  possible  the  recovery  of  much  pillar 
coal,  otherwise  irremovable,  under  townsites.  Less  liability  to  accidents 
from  roof  falls  is  another  advantage  of  flushing.  Example  60  illustrates 
the  first  application  of  flushing  to  metal  mining,  and  it  is  already  having 
many  imitators  in  the  Rand.  The  flushing  system  is  applicable  to  all 
metal  mines  where  the  ore  lies  in  beds  of  large  area,  and  where  plenty  of 
tailing,  or  sand,  and  water  is  available.  Its  application,  in  any  given 
case,  is  a  commercial  question;  the  cost  of  digging,  transporting,  and 
damming  the  filling  and  of  repumping  the  flushing  water  being  set  off 
against  the  net  value  of  the  recovered  pillars  and  surface  acreage. 


CHAPTER  XXIII 
PRINCIPLES  OF  MINE  EVALUATION 

Mineral  deposits,  especially  those  of  the  rarer  metals,  are  often 
irregular  in  size  and  uncertain  in  tenor  and  these  peculiarities  introduce 
the  element  of  chance  into  their  extraction.  Though  mining  is  thus 
naturally  more  or  less  speculative,  it  is  by  no  means  "just  a  gamble"  as 
the  average  layman  is  apt  to  opine,  for  sound  finance  with  the  aid  of 
applied  science  has  in  recent  years  gone  far  to  place  mining  upon  as  legiti- 
mate an  investment  basis  as  agriculture  or  transportation. 

In  evaluating  a  mine  we  have  two  forms  of  property  to  consider, 
land-values  and  improvements.  Of  the  two  forms  the  first  is  paramount, 
for  with  a  just  sufficient  quantity  S  of  "ore"  (payable  metallic  or  non- 
metallic  mineral)  in  the  mine  to  return,  with  interest,  the  net  expenditure 
on  improvements  necessary  for  its  extraction,  the  property  has  no  value 
at  all  as  mineral  land.  Any  excess  of  ore  beyond  S  grants  a  land-value 
to  the  mine,  while  any  deficit  under  S  means  a  loss  of  liquid  capital  in  the 
development. 

It  is  thus  essential  to  observe  and  study  the  physical  features  of  the 
property,  as  the  basis  for  evaluation,  and  these  may  be  grouped  under 
twelve  topics  as  follows:  (1)  Quality  of  ore;  (2)  Quantity  of  ore;  (3)  Loca- 
tion for  transportation;  (4)  Available  markets;  (5)  Local  topography;  (6) 
Water  conditions;  (7)  Climate;  (8)  Food  supply;  (9)  Fuel  supply;  (10)  Labor 
conditions;  (11)  Government;  (12)  Necessary  investment. 

Topic  (1)  is  got  by  first  sampling  the  ore  body;  and  then  making 
the  chemical  analyses  and  working  tests  on  these  samples  necessary  for 
determining  the  values  in  .the  ore  and  their  commercial  utilization. 
Topic  (2)  may  be  estimated  from  the  development  openings  and  from 
the  geological  nature  of  the  deposit. 

On  this  topic  G.  E.  Collins  has  presented  as  a  practical  example,  a 
section  (Fig.  146)  of  a  fissure-vein,  the  ore  reserves  of  which  he  classifies 
as  "positive"  (proved  on  three  sides),  "probable"  (proved  on  two  sides), 
and  "possible"  (proved  on  one  side).  Upon  the  case  thus  supposed,  he 
thus  remarks. 

"Most  of  us  have  known,  many  of  us  have  experienced,  cases  where 
blocks  of  ore  so  exposed  (even)  on  four  sides  have  been  found  to  enclose 
a  large  barren  patch  in  the  center.  The  amount  of  weight  to  be  attached 
to  exposure  on  four  sides,  on  three,  on  two,  or  on  one  only,  varies  with  the 
conditions  of  each  particular  case  .  .  .  Who,  after  examining  the 

294 


PRINCIPLES   OF   MINE   EVALUATION 


295 


section  (Fig.  146),  and  bearing  in  mind  the  phenomenal  lateral  extension 
of  the  ore-body  shown,  can  doubt,  that  far  more  reliance  can  be  placed 
on  blocks  marked  in  the  section  as  '  'positive  ore,"  even  although  some  of 
them  are  exposed  on  three  sides  only,  than  on  many  blocks  opened  up  on 
four  sides  in  veins  where  the  ore-bodies  are  of  the  bunchy  and  erratic 
type  which  we  must  recognize  to  be,  after  all,  by  far  the  most  frequently 
encountered?  Even  the  "possible  ore"  of  Fig.  146 — ground  where  the 
ore  is  proved  to  exist  on  one  side  only — can  be  depended  on  to  a  far 
greater  extent  than  usual. 

1  'My  contention  is  that  the  amount  of  evidence  to  be  required  when 
making  estimates  of  ore  tonnage,  and  the  number  and  nature  of  classes 


"Positive  Ore"  -  Ore  ground  proved  on  3  sides 

"Probable  Ore"-  Ore  ground  proved  on  2  sides— - 

'•'Possible  Ore"-  Ore  ground  proved  on  1  side 

Ore  ground  stoped  out 

1  inch  =  890  feet 

FIG.  146. 

into  which  estimates  should  be  divided,  must  depend  on  our  general 
conclusions  as  to  the  nature  and  permanence  of  the  ore-bodies  in 
the  mines  under  consideration,  the  distance  between  the  workings,  etc. 
I  do  not  think  that  any  rules  can  be  made  which  will  lessen  the  necessity 
for  dependence  on  the  examining  engineer's  individual  judgment;  and  I 
distrust  all  cast-iron  classifications,  which  do  not  allow  for  the  infinite 
complexity  of  natural  conditions.  The  only  useful  general  rule  is  that  no 
estimate  of  tonnage  should  be  made  unless  accompanied  by  sketches 
indicating  the  basis  on  which  it  rests." 

Topic  (3)  relates  to  the  transportation  facilities  such  as  existing 
water  or  land-ways  or  the  cost  of  constructing  same  to  reach  the  nearest 
navigable  water  or  railroad  leading  to  market;  also  to  the  animal  or  other 
power  available  for  transportation.  This  topic  determines  the  cost  of 
shipping  the  output  to  market  and  of  importing  those  supplies  not 
obtainable  locally.  The  available  markets  of  (4)  may  be  fluctuating  or 
steady  in  their  demands  and  may  accept  in  some  cases  the  crude,  in  others 
only  the  finished  mineral  product.  The  local  topography  of  (5)  not  only 
fixes  the  possible  methods  of  mine  development,  but  also  affects  the  loca- 


296  MINING    WITHOUT    TIMBER 

tion  of  works,  housing  and  transportation.  Under  (6),  or  water  condi- 
tions, we  consider  water  power,  mine  drainage,  water  supply  for  works 
and  people,  and  the  condition  of  roads  and  waterways. 

The  climate  (7)  of  the  past  has  affected  the  oxidation  or  enrichment 
of  the  ore-body  and  at  present  it  affects  wind  power,  health,  food  supply 
and  the  conditions  of  outdoor  work.  Topic  (8)  of  food  supply  depends 
on  the  regions'  soil  and  climate  and  on  the  skill,  number,  and  industry  of 
the  present,  or  possible  future,  local  farmers. 

The  fuel  supply  of  (9)  is  a  question  of  the  existence  near  the  mine  of 
available  forests,  peat  bogs,  oil  or  gas  wells,  or  coal  seams.  Topic  (10) 
of  labor  hangs  entirely  on  the  local  population  in  those  many  cases  where 
induced  immigration  is  impractical  except  for  the  official  staff.  The  rate 
of  wages,  the  adaption  to  mining  work,  the  industry  and  the  docility  of  the 
natives  are  all  vital  practical  questions.  The  government  of  (11)  is  often 
the  determining  factor  for  an  enterprise,  not  only  does  its  system  of 
taxation  more  or  less  affect  costs  but  revolutions  or  predatory  officials 
may  sometimes  render  inadvisable  any  investment  whatever. 

The  capital  necessary  for  (12)  can  only  be  calculated  after  a  summa- 
tion and  correlation  of  the  previous  eleven  topics.  The  total  investment 
will  be  apportioned  between  "fixed"  and  "working"  capital;  the  former 
being  the  amount  needful  to  expend  in  mine  development  and  equip- 
ment and  in  the  construction  of  works  and  roads,  to  make  the  property 
suitably  productive,  and  the  latter  being  the  additional  sum  required 
during  productive  operations. 

The  following  is  an  attempt  to  form  a  formula  by  which  a  mine  can  be 
quickly  evaluated,  after  the  above  pertinent  physical  data  have  been 
collected  from  observations  on  the  ground  by  a  competent  mining  engineer. 

ASSUMPTIONS 

Let    G  =  price  to  be  paid  for  the  mining  property. 

Let    M  =  cost  of  developing  the  mining  property  to  yield  y  tons 

of  ore  daily. 

Let    P  =  cost  of  suitable  plant  to  treat  y  tons  of  ore  daily. 
Let     p  =  value  of  said  plant  when  the  mine  has  been  exhausted. 
Let     C  =  total  fixed  capital  investment. 
Let  W  =  working  capital  investment. 
Then  let  C  +  W  =  total  capital  investment. 

Let     y  =  required  yield  of  ore  daily  in  tons. 

Let    Y  =  yield  of  ore  yearly  in  tons. 

Let     d  =  number  of  producing  days  per  year. 

Let     u  =  average  operating  profit  per  ton  of  ore. 

Let    R  =  rate  of  interest  to  be  earned  on  total  investment  of 

(C  +  W). 
Let     r  =  rate  of  interest  to  be  earned  on  sinking  fund  annuity. 


PRINCIPLES    OF   MINE   EVALUATIONS  297 

Let     v  =  tons  of  positive  ore  available  in  mine  (see  Fig.  146). 

Let     x  =  tons  of  probable  ore  available  in  mine  (see  Fig.  146). 

Let     z=tons  of  possible  ore  available  in  mine  (see  Fig.  146). 

Let    m  =  fractional  factor  to  change  probable  to  positive  ore. 

Let     n  =  fractional  factor  to  change  possible  to  positive  ore. 

Let    Q  =  tons  of  total  ore  available  in  mine. 

Let      t  =  time  in  years  to  exhaust  mine  at  rate  Y. 

Let     b  =  time  in  years  for  mine  to  reach  production  of  y  tons. 

Let    A  =  annuity  to  be  paid  to  sinking  fund  to  equal  C  at  end  of 

t  years  at  r,  compound  interest. 
Let     a  =  annuity  to  be  paid  to  sinking  fund  to  equal  C  at  end  of 

t  years  at,  r,  simple  interest. 

DERIVATION  OF  FORMULAE 

We  have   directly  from  the  foregoing   assumptions  the  following 
equations : 

C  =  G  +  M+P  (1). 

Y  =  dy  (2). 

Q  =  v  +  mx  +  nz  (3). 

Q  =  tY  (4). 

Then  since  we  must  balance  the  two  debits,  of  the  interest  to  be  paid 
during  (t+b)  years  and  the  sinking-fund  annuity  to  be  paid  during  t 
years,  against  the  two  credits,  of  the  operating  profit  from  the  total 
available  ore  and  the  selling  value  of  the  abandoned  plant,  we  have: 

(t  +  b)(C  +  W)  R  +  tA  =  tYu  +  p  (5). 

To  determine  A,  we  can  substitute  in  that  algebraic  formula  for  the 
annuity  which  involves  the  rate  of  compound  interest,  the  number  of 
annuity  payments  and  the  capital  to  be  refunded  and  get: 


Substituting  in  (5)  this  value  of  A,  we  have: 

(t  +  b)(C  +  W)R  +  7~^r-   -  =  tYu+p  (7). 


With  the  observed  data,  and  equations  (1),  (2),  (3),  and  (7),  we  can  pro- 
ceed to  evaluate  the  mine.  Equation  (7)  includes  nine  factors,  any  one 
of  which,  except  t,  can  be  determined  by  solving  a  simple  equation. 
When  t  is  unknown  (as  is  commonly  the  case) ,  it  is  difficult  to  solve  the 
equation  by  common  algebraic  methods.  To  obviate  this  difficulty  it 
may  be  assumed  that  simple,  instead  of  compound,  interest  is  to  be  earned 
on  the  sinking-fund  annuity;  and  the  resulting  difference  in  the  result 
will  act  as  a  safeguard  on  the  side  of  the  mine-buyer.  Additional  safe- 
guards for  the  buyer  will  be  the  assumption,  in  equation  (7),  that  the 


298  MINING    WITHOUT    TIMBER 

factors  C  and  W  are  invested  at  once,  whereas  their  expenditure  usually 
occupies  a  considerable  period. 

With  simple  interest,  the  sinking-fund  annuity  (a)  becomes  the  first 
term  of  an  arithmetical  progression,  of  which  the  common  difference  is  the 
annual  interest  (ar)  gained  on  the  annuity,  the  sum  is  the  capital  (C)  to 
be  repaid,  and  the  number  of  terms  is  (t)  the  period  of  years.  Then,  by 
substitution  in  that  algebraic  formula  for  the  sum  which  involves  the 
first  term,  the  common  difference  and  the  number  of  terms  we  have: 

20 

TTx  (8). 


Substituting  this  value  of  a  for  A  in  (5),  we  have: 

20 

(t  +  b)  (C  +  W)  R  +  - — — —  =tYu  +  p  (9). 

To  determine  from  equation  (9)  the  unknown  factor  t,  will  involve 
only  the  solution  of  a  quadratic  equation;  and  this  is  most  conveniently 
performed  after  the  substitution  of  the  numerical  values  of  the  other 
factors;  since  the  direct  solution  for  t,  of  (9)  as  a  literal  equation,  is  very 
lengthy. 

If  no  interest  be  earned  on  the  annuity  payments,  the  second  term  of 
equation  (9)  will  become  equal  to  C,  and  we  shall  have: 

(t  +  b)  (C  +  W)  R  +  C  =  tYu  +  p  (10). 

Since  equation  (10)  involves  only  first  powers,  it  can  be  used  as  a 
quick  check  on  the  approximate  accuracy  of  the  solution  of  equation  (9) 
fort. 

PRACTICAL  EXAMPLES 
I 

On  a  certain  property  in  the  southwest,  examined  by  the  author,  it 
was  required  to  find  the  minimum  available  ore  that  must  be  found  by 
prospecting  operations  to  warrant  the  capital  expenditure  required  to 
inaugurate  production  on  a  given  scale.  The  unknown  factors  were 
hence  Q  and  t,  and  the  known  were  estimated  from  the  collected  data 
to  be: 

G  =  $40,000;  M  =  $56,000;  P  =  $39,000;  p  =  $10,000;  W  =  $15,000; 
y  =  100  tons;  d=300  days;  u  =  $3;  R  =  0.15;  r  =  0.06;  b  =  2  years. 

From  equation  (1),  0  =  40,000 +  56,000 +  39,000  =  $135,000. 

From  equation  (2),  Y  =  300X100  =  30,000  tons. 

From    equation     (9),     (t +  2  (135,000  +  15,000)  0.15+  -^^^— 

2  +  0.06(t-l) 
=  t(30,000x3) +10,000;  or 


PRINCIPLES    OF    MINE    EVALUATION  299 


27t(1.94  +  0.06t)-  14(1.94  +  0.061)  -108  =  0,  or 
0.81t2  +  25.77t-67.58  =  0;  whence 
t  =  2.42  or-  34.24. 

It  is  evident  that  the  positive  value  of  2.42  years  is  the  one  desired. 
Substituting  the  values  of  t  and  Y  in  equation  (4),  we  have: 

Q  =  2.42X30,000  =  72,600, 

the  number  of  tons  of  available  ore,  that  should  be  found  by  prospecting, 
to  satisfy  the  conditions  of  the  case. 

II 

A  consumer,  using  5000  tons  of  a  certain  metal  yearly,  wishes  to 
acquire  a  mine  which  would  furnish  his  whole  supply.  He  has  found  a 
mine  which,  by  the  expenditure,  besides  the  purchase-price,  of  $200,000 
for  development  and  plant,  and  the  provision  of  $50,000  working  capital, 
would  enable  him  to  obtain  annually  the  required  supply  of  metal  from 
60,000  tons  of  its  ore,  at  an  operating  profit  of  $4  per  ton.  A  year  will  be 
required  to  develop  and  equip  the  property,  and  the  available  ore  will 
last,  at  the  required  rate  of  production,  for  twenty  years,  at  the  end  of 
which  period  the  plant  will  be  worthless.  If  interest  on  the  total  invest- 
ment be  reckoned  at  6  per  cent,  and  on  the  sinking-fund  annuity  at  5  per 
cent.,  and  if,  in  addition,  it  is  necessary,  while  working  the  mine,  to  make 
a  net  saving  of  1  cent  a  pound  on  the  whole  metal  supply  of  the  consumer, 
what  price  could  he  afford  to  pay  for  the  mine? 

Here  C  and  G  are  the  unknown  factors,  and  we  have: 
M+P  =  $200,000;  W  =  $50,000;  u  =  $4;  p  =  0;  annual  saving  on  metal, 
$100,000;  Y  =  60,000  tons;  R  =  0.06;  r  =  0.05;  t  =  20  years;  and  b  =  l  year. 

From  equation  (1),  C  =  G  +  200,000. 

Substituting  this  value  in  equation  (7),  and  remembering  that  the 
sinking-fund  annuity  is  to  be  increased  by  the  $100,000  of  annual  saving 
on  metal  supply,  we  have: 


(20  +  1)  (0  +  200,000  +  50,000)  QM+--  ^  +  (20X100,000) 

(20X60,OOOX4)+0;  or 


nnn 
1.  26  (G  +  250,000)+  —  -----  '-      -+2,000,000  =  4,800,000;  or 

2.084(G  +  250,000)  +G  +  200,000  +  1.654  (2,000,000) 
=  1.654  (4,800,000);  or 

G  =  $1,264,500, 
the  maximum  allowable  price  for  the  mine. 


300  MINING  WITHOUT  TIMBER 

CONCLUSIONS 

The  practical  value  of  such  calculations  as  the  foregoing  may  be 
plausibly  questioned,  at  least  as  regards  all  mines  other  than  collieries,  for 
which  the  quantity  of  available  reserves  can  be  estimated  with  a  degree 
of  confidence  and  precision  not  usually  attainable  in  mines  of  the  metallic 
ores.  In  reply  to  this  probable  criticism  the  author  begs  to  offer  the 
following  observations: 

1.  There  are,  in  fact,  besides  collieries,  more  mines  than  we  commonly 
realize,  the  actual  reserves  of  which  can  be  measured,  and  the  probable  or 
potential  reserves  estimated.     Among  such  might  be  instanced  many 
quarries,  massive  ore-bodies  already  explored  by  boring,  etc.     To  all 
such  cases,  mathematical  formulas  of  valuation  are  directly  applicable. 

2.  With  regard  to  the  very  large  number  of  metal-mines,  in  which  v, 
the  certain  ore-reserve,  is  relatively  small,  while  x,  the  probable,  and  z, 
the  possible  reserve,  are  so  uncertain  as  to  make  the  determination  of  m 
and  n,  as  the  moduli  representing  reasonable  expectation,  merely  a 
function  of  the  temperament  of  the  observer,  yet  even  in  such  cases,  the 
observer  himself  may,  very  likely,  be  steadied  in  his  judgment,  and  aided 
to  form  prudent  conclusions,  by  such  calculations  as  will  clearly  show 
him  to  what  .extent  his  hopes,  fears  and  guesses  enter  into  his  opinions. 
It  may  seem  absurd  to  employ  mathematical  methods  in  the  discussion  of 
data  so  largely  indefinite;  but  the  quantitative  determination,  even  of 
our  ignorance,  is  a  recognized  application  of  mathematics;  and  "the 
probable  error"  has  a  value  of  its  own,  not  less  real,  though  it  be  less 
authoritative,  than  the  rigorously  demonstrated  certainty.     In  other 
words,  the  discussion  by  exact  methods,  even  of  more  or  less  uncertain 
data,  is  a  valuable  check  upon  the  hasty,  sentimental  or  temperamental 
general  impressions  which  often  claim  the  authority  of    conclusions. 
Such  a  check  is  the  more  important,  because  many  mining  investors  are 
tempted  to  overlook  the  essential  proposition  that  a  mine  is  l  'a  candle, 
burning  as  both  ends";  that  its  value  is  constantly  diminished  by  its 
product;  and  that  its  annual  profits  must  cover,  not  only  a  satisfactory 
income  from  the  investment  which  has  been  made  in  it,  but  also  the  pro- 
gressive  repayment   of   the   investment   itself.     All   this   sounds   very 
elementary;  yet  it  is  too  often  overlooked  in  the  enthusiasm  of  specula- 
tion.    Each  investor,  if  he  thinks  of  it  at  all,  dismisses  it  from  his  mind 
with  the  reflection  that  he  will  have  abundant  opportunity  to  "unload" 
during  the  period  of  dazzling  prosperity  which  he  foresees  for  the  mine. 
This  is  the  chief  reason  why  mining  has  not  yet  become  universally  a 
regular  business.     That  desirable  result  will  be  greatly  promoted  when 
investors  purchase  a  mine,  either  with  the  positive  intention,  not  of 
selling  it  out,  but  of  working  it  out,  or  else,  at  a  price  based  upon  the 
hypothesis  of  such  an  intention.     Consequently,  the  more  this  considera- 
tion is  emphasized  by  theoretical  calculations  of  value  like  the  preceding, 
the  better. 


APPENDIX  I 


ARTICLES  FROM  WHICH  MUCH  OF  THE  SUBJECT-MATTER  OF  THE  BOOK  WAS 

EXCERPTED 


Book 

Published  Article 

Chap- 

Ex- 

ter 

imnle 

Title 

Magazine 

Date 

Writer 

1                            Exnloaivps                      1    Mininir  &-  "Flno1   WnrlH 

Apr.  29,  May  13 

The  author 

1911  

2 

Blasting 

Mining  &  Eng.  World 

Mar.  11,  1911    .  .  . 

The  author 

3 

Compressed  Air  

Eng     &   Min.    Journal 

Jan.  18,  1908  

The  author 

4 

Excavations 

Mining  &  Eng.  World  . 

May  27,  1911 

The  author 

5 

Drainage  

Mining  &  Eng.  World.  . 

June  22    1911  ... 

The  author 

Ditto  

Nov.  28,  1908 

E.  B.  Kirby 

6 

1 

Cuba  

Trans  Min   Eng  

March,  1911  

D.  Woodbridge 

2 

Mesabi.  j   Mining  Science  Dec.  24-31,  1908. 

The  author 

3 

Utah  Copper  

Mines  &  Methods.    Sept.,  1909  

C.  T.  Rice 

4 

Nevada-Con  j 

Mines  &  Methods  
Min.  &  Sc.  Press  

Sept.,  1909  
Mar.  4,  1911  

C.  T.  Rice 
E.  E.  Barker 

5 

Danville  

Mines  &  Minerals     .... 

Oct.,  1907,  Sept., 

Anonymous 

1910 

7 

6 

Puertocitos  

Mines  &  Methods  

Feb.,  1910  

C.  T.  Rice 

7 

Mesabi.  . 

Ditto,  Example  2 

8 

Traders 

Ditto,  Example  2  

Apr.  22,  1909  The  author 

9 

Alaska  Tread  well.  . 

Trans.  Min.  Eng  

Vol.  34  R.  A.  Kinzie 

8 

10        Southeast  Mo  

Mines  &  Minerals  

Nov.andDec.,190l|   The  author 

Eng.  <fc  Min.  Journal.  .  . 

Feb.  12,  1910  

Anonymous 

11        Southwest  Mo  Mines  &  Minerals  

Nov.,  1907  

J.  H.  Polhemus 

Mining  &  Eng.  World.  . 

Sept.  26,  1908  ...     Otto  Ruhl 

12 

Bisbee  Eng.  &  Min.  Journal.  .  . 

July  23,  1910  M.  J.  Elsing 

13 

Section  21  Mining  Science  

Jan.  7,  1909  The  author 

9 

14        Wolverine  Mining  &  Eng.  World  .  . 

Mar.  26,  1910  The  author 

15     '   Homestake  

Eng.  &  Min.  Journal.  .  . 

July  9,  1910  J.  Tyssowski 

16     !  Gratz  

Eng.  &  Min.  Journal.  .  . 

Apr.  6,  1907  

The  author 

17     i   Alaska-Tread  well.  .  . 

Ditto,  Example  9. 

18         Veta  Grande  

Eng.  &  Min.  Journal  .  . 

Nov.  12,  1910  .... 

M.  J.  Elsing 

10 

19 

South  Range  

Mining  &  Eng.  World  .  . 

Mar.  26,  1910  

The  author 

20 

Minnesota  

Mining  Science  .  .    

Dec.  17,  1908  

The  author 

21 

Superior  &  Boston.  . 

Mines  &  Minerals  

Sept.,  1910  

R.  L.  Herrick 

22         Metcalf  

Eng.  &  Min.  Journal.  .  . 

July  16,  1910  

P.  B.  Scotland 

23 

Bisbee  

Mines  &  Minerals  

Feb.  1907  

The  author 

11 

24 

Los  Pilares  

Mines  &  Minerals  

Sept.,  1910  '  E.  M.  Robb,  Jr. 

25        West  Australia  

Mines  &  Methods  

Dec.,  1910  R.  Allen 

26         British,  

Mines  &  Minerals.  .  May,  1907  1  A.  J.  Moore 

27        Proprietary 

Mines  &  Minerals  Mn,v.  1907  A.  J.  Moore 

12 

28        Central  

Mines  &  Minerals  

May,  1907  A.J.Moore 

29     '   King  

Ditto,  Example  22. 

30         Coronado  

Ditto,  Example  22. 

31         Los  Pilares  

Ditto,  Example  24. 

13 

32        Miami. 

Mines  &  Minerals                 Tnlv    191O 

R.  L.  Herrick 

33         Boston-Con  

Mines  &  Methods  

Sept.,  1910  

C.  T.  Rice 

34        Duluth  

DLto,  Example  18. 

301 


302 


APPENDIX   I 


Book 


Published  Articles 


Chap- 
ter 

Ex- 
ample 

Title 

Magazine 

Date 

Writer 

14 

35 

Hartford 

Ditto  Example  13 

36 

Pi;>neer  ... 

Minin1"  Science  .  .  . 

Dec.  10,  1908..   .  . 

The  author 

37 

Utah-Copper. 

Ditto,  Examole  3 

15 

38 

Pewabic  ... 

Minin<*  Science  

Apr.  29,  1909  

The  author 

39 

Mowry  

Mines  &  Minerals  .... 

Jnly,  1907  

The  author 

40 
41 

Detroit-Copper  
Comnieroial. 

Min.  &  Sci.  Press  
Mines  &  Minerals  .... 

Dec.  24,  1910  
Oct.,  1907  

W.  L.  Tovote 
The  author 

42 

Inspiration  . 

Mines  &  Methods 

June,  1909  

C.  T.  Rice 

16 

17 

18 
19 

43 
44 
45 
46 

47 

48 

50 
51 
52 

Old  Jordan  
Cumberland-Ely  
Oversight  
Lake  Superior  

Mercur  
Kimberley  

Roof-Pressure  
Montour  
Providence  
Vintondale 

Mines  &  Minerals  
Mines  &  Methods  
Ditto,  Exapmle  6. 
Mining  Science  

En?.  &  Min.  Journal.  .  . 
"Diamond      Mines      of 
S.  A." 
Mines  &  Minerals  
Trans.  Min.  Eng  
Mines  &  Minerals  
Coal  Mining  Inst   .  . 

Oct.,  1907  
Sept.,  1909  

Feb.   18  and  Apr. 
22,  1909. 
June  18,  1910.... 

April,  1907  
October,  1891  
April,  1907  
June,  1907  

The  author 
C.  T.  Rice 

The  author 

R.  H.  Allen 
G.  F.  Williams 

H.  Briggs 
H.  H.  Stoeck 
Anonymous 
J.  I.  Thomas 

20 

53 
54 
55 

Nova  Scitia  
Gen.  Pillar  
Retreatinf 

Mines  &  Minerals  
Mines  &  Minerals  
E.  &  M.  Journal  

June,  1909  
Jan.,  1899  
Dec   26,  1908   .  .  . 

J.  G    MacKenzie 
J   T.  Beard; 
H.  J.  Nelms 

21 
23 

56 
57 

58 

60 

Adv.  Retreat  
Connellsville  
Pittsburg  

Rand  
Evaluation  

E.  &  M.  Journal  
Mines  &  Minerals  
Traps.  Min.  Eng  

E.  &  M.  Journal  
E.  &  M.  Journal  
Trans.  Min.  Eng  

Apr.  23,  1910  
July,  1907  
March,  1910  

Feb.  5,  1910  
March  14,  1903.  .. 
Apr.   1907  

H.  J.  Nelms 
G.  S.  Baton 
F.Z.Schellenberg 

Anonymous 
G.  E.  Collins 
The  author 

INDEX 


Aliabatic.    See  Air. 

Adits,  58-59,  221 

Advancing;  system  for  seams,  229,  238, 

242,  245,  250,  255,  260,  292 
Advancing-retreating  system  for  seams, 

230,  266,  272 
Africa,  41,  224,  229,  283 
Air,  compressed,  29-32,  98,  169,  184 
drill.    See  Drill. 
fresh.     See  Ventilation. 
Alaska  Tread  well  mine,  Alaska,  88,  113, 

286 

Alf tofts  colliery,  Eng.,  233 
America  mine,  Mex.,  210 
American  Z.  L.  and  S.  Co.,  Mo.,  95 
Amvis,  16 
Analysis,  71,  294 
Anthracite  district,   Pa.,  43,  55,  60,  82, 

276 

Antoine  Ore  Co.,  Mich.,  87 
Appalachian  beds,  229 
Arching  system,  119 
Arizona  Copper  Co.,  158 
Atlantic  mine,  Mich.,  118,  121 
Augers,  62,  206,  209,  212,  220 
Australia,  143,  145,  155,  287 


B 


Back-caving.     See  Caving. 
Bacon,  Roger,  7 
Baldwin  locomotives,  141 
Baltic  mine,  Mich.,  118 
Barrier  Range,  N.  S.  W.,  145 
Barriers.     See  Dams. 
Bellefontaine  Machine  Co.,  O.,  83 
Bingham,  Utah,  53,  72,  76,  171,  187, 

206 

Bisbee,  Ariz.,  50,  100,  130 
Biwabik  mine,  Minn.,  67 
Black  Diamond  mine,  Pa.,  282 
Black  Hills,  S.  Dak.,  108 
Blasting,  calculations,  18 


Blasting,  massive  rock,  20,  74,  81,  116, 
129,  173,  189 

stratified  rock,  21-24,  83,  85,  94,  97 

timber,  206,  209,  212,  214 

seams,  26,  253 

See  also  Chambering,  Electric,  Ex- 
plosives, Multicharging,  Tunnel. 
Block-caving.     See  Caving. 
Block  10  mine,  N.  S.  W.,  149 
Bobbinite,  8 

Bonne  Terre,  Mo.,  91,  93,  94 
Boston  and  Montana  mine,  Nev.,  77 
Boston  Consolidated  mine,   Utah,    171, 

207,  288 
Boyle's  law,  5 
Brattices,  263,  267 
Breaker.     See  Coal. 
Breast-stoping,  26,  93,  96,  122,  152,  206, 

208,  219,  243,  290 
Brickyard  mine,  Utah,  220 
British  mine,  N.  S.  W.,  145,  149,  151 
Broken  Hill,  N.  S.  W.,  145,  152,  155 
Brushing.     See  Roof. 

Bucyrus  shovel,  68,  75,  81 
Bulldozing,  74,  81,  88,  90,  114 
Bull's  Head  colliery,  Pa.,  245 
Buggies  mine,  247,  262,  292 
Bureau  County,  111.,  238 
Butte,  Mont.,  26,  31 


Cages,  184,  195,  219 

Calumet  and  Arizona  mine,  Ariz.,  100 

and  Hecla  mine,  Mich.,  54 
Calumet  and  Pittsburg  mine,  Ariz.,  50 
Cananea,  Mex.,  84,  115,  176,  210 
200,      Canisteo,  mine,  Mian.,  69 
Caps,  blasting,  2,  15 
Capote  mine,  Mex.,  210 
Carbonite,  16 
Cars,  coal,  240,  248,  259,  264 

ore,  61,  88,  97,  120,  133,  141,  170, 
171,  187,  189,  195,  209,  218,  219 

303 


304 


INDEX 


Cars,  surface,  67,  75,  81 
Cartridge,  3 ' 

Caving,  back,  104,  158,  181,  183,  187,  214, 
221,  224,  258,  288,  289,  291 

block,  170,  174,  178,  192,  195,  196, 
200,  289,  291 

chute,  181,  183,  187,  289 

formation.     See  Roof  and  Seams. 

hangwall,  125 

loss  in,  190,  206,  216,  223,  289,  290, 
291 

pillar,  165,  288,  291 

room,  104,  170,  214,  222,  227,  290, 

291 

Central  mine,  N.  S.  W.,  155 
Chambering,  25,  69,  73,  81,  85,  86,  96, 

122,  164 
Champion  mine,  Mich.,  119 

powder,  81 

Chapin  mine,  Mich.,  57,  217 
Charles'  law,  5 
Chocks.     See  Cribs. 
Chute-caving.     See  Caving. 
Chutes,  loading,  87,  182,  188,  214,  244, 

262,  289 

Cleaning.     See  Coal  and  Sorting. 
Cleavage,  25,  231,  250,  269 
Clod,  35 
Coal  breaker,  276 

commission,  235,  276 

cutting,  Hand,  235,  240,  248,  252,  266 
machine,  26,  232,  253,  258,  266, 
272,  275 

mining,  26,  35,  41,  43,  82,  228-242, 
245-283 

washery.    See  also  Seams,  and  Costs. 
Cogs.     See  Cribs. 
Colliery.     See  Coal. 
Colliery-steelite,  16 
Collins,  G.  E.,  294 
Commercial  mine,  Utah,  200 
Comparison  systems,  285-293 
Compression.     See  Air. 
Comstock  lode,  Nev.,  54,  59,  108 
Consolidated  Mercur  mines,  Utah,  220 
Connellsville  district,  Pa.,  268 
Continental  mine,  Pa.,  271. 
Continuous-face.     See  Longwall. 
Conveyors,  83,  250,  278,  292 
Copper  Flat  mine,  Nev.,  77 

Queen  mines,  Ariz.,  130 
Coronado  mine,  Ariz.,  158 
Costs,  drilling,  63,  71,  77 


Costs,  haulage,  84 

mining  ore-bodies,  97,  99,  107,  111, 
117,    120,    145,    149,   170,   174, 
180,  183,  187,  191,  193,  196,  198, 
205,  207,  209,  210,  218,  224 
seams,  244,  254 
surface,  85,  87 
steam-shoveling,  66,  68,  76,  82 

Cranes,  82,  98A 

Creep,  40,  241,  264 

Cribs,  146,  241,  248,  254,  256,  287,  288, 
292 

Cripple  Creek,  Colo.,  28,  59 

Cross-cut  system,  152,  287 

Cuba,  62,  285 

Culm,  276,  278,  293 

Cumberland-Ely  mines,  Nev.,  207 

Cyanide  tails,  143,  283,  284 

D 

Dams,  artificial,  55,  61 
deposition,  58A 
natural',  56,  280,  281,  283 
Danville  colliery,  111.,  83 

mines,  Pa.,  242 
Davey  mines,  Mo.,  95 
Delaware  and  Hudson  Coal  Co.,  Pa.,  276 
D.  L.  and  W.  Coal  Co.,  Pa.,  276 
Denver  Engineering  Works,  Colo.,  206 
Desloge  mine,  Mo.,  93 
Detroit  mine,  Ariz.,  196 
Development,  mine,  20 
Dip-working  of  seams,  236 
Dodson  colliery,  Pa.,  276,  278,  281,  282, 

283 

Doe  Run  mine,  Mo.,  91 
Dor  ranee  colliery,  Pa.,  280,  281,  282 
Douglas  Island,  Alaska,  88,  113 
Drainage,  calculating,  47 
massive  rock,  53 
stratified  rock,  52,  230,  242,  256 
unconsolidated    rock,    47,    67.     See 

also  Adits,  Pumps,  and  Water. 
Dressing.     See  Sorting. 
Drill,  air-hammer,  21,  24,  95,  111,  161, 

172,  173,  196 

air-piston,  74,  86,  88,  90,  94,  96,  102, 
107,  111,  113,  114,  169,  171,  183, 
184,  187,  189,  195,  200,  212,  288 
churn,  hand,  62,  69,  85 

power,  blasting,  26,  73,  80,  164 
prospecting,  70,  77,  92,  230,  279 
diamond,  71,  91,  122,  183 


INDEX 


305 


i      Drill,  jumper,  84,  220,  223,  227 

-pointing,  20-27 
Drummond  colliery,  N.  S.,  255 

system,  201 
Drywalling,  107,  119,  236,  241,  243,  246, 

248,  280,  283,  287 
Ducktown,  Tenn.,  95 
Duluth  mine,  Hex.,  176,  210,  288 
Dupont  powder,  81 
Dust,  26,  230 
Dynamites,  13.     See  also  Blasting. 

E 

East  Nome  mine,  Mich.,  214,  219 
Electric  blasting,  19,  28,  74,  81,  85,  196, 
209 

exploders,  2 

power,  252.     See  also  Haulage. 
Ely,  Minn.,  183 
Ely,  Nev.,  76,  207,  285 
England,  41,  233,  235,  250 
English  Iron  Works,  Mo.,  97 
Excavations,  44-46.     See  also  Roof. 
Excavator,  dray-line,  64,  87,  285 
European  mines,  229,  255,  276,  293 
Eveleth,  Minn.,  216 
Explosions,  mine,  230,  264 
Explosives,  calculating,  4 

chemical,  8 

detonating,  2,  74,  81 

igniting,  1,  173,  206 

loading,  3 

mechanical,  7 

-chemical,  12.     See  also  Blast- 
ing, Electric,  Powder. 


Fayal  mine,  Minn.,  67,  105,  215,  219 

Fayal  rules,  34,  41 

Filling  ore-bodies,  119,  123,  143,  147,  155, 

159,  164,  207,  218,  287,  288 
seams,  241,  244/248,  254,  266,  275, 

283,  293 

Flat  River,  Mo.,  91 
Flowage  rock,  46 
Forcite,  14 
Fracture.     See  Roof. 
Frick  Coal  Co.,  Pa.,  272 
Flusing  pipe,  279 

slush,  278,  283,  284 

system,    276,    283,    293.     See    also 

Filling  and  Roof. 
Fulminates,  11,  14 
20 


Fuse,  2.     See  also  Blasting. 


Garden  City  Fan  Co.,  111.,  108 

Gas,  230,  242,  257,  281,  291 

Gelignite,  14 

General  Electric  locomotive,  141 

Geneva,  N.  Y.,  49 

Glenn  mine,  Minn.,  215,  217,  220 

Globe  district,  Ariz.,  124,  165,  201 

Glory-hole.     See  Milling. 

Gobbing.     See  Filling. 

Gogebic  Range,  Mien.,  184,  214 

Golden  Gate  mine,  Utah,  220 

Goodman  locomotive,  184 

Graham  County,  Ariz.,  128,  156,  158,  196 

Gratz  mine,  Ky.,  112 

Grundy  County,  111.,  238,  291 

Gulf  of  Mexico,  49 

Guncotton,  12 

H 

Hamilton  mine,  Mich.,  217 
Hangwall-caving.     See  Caving. 
Hartford  mine,  Mich.,  181,  289 
Haulage,  air,  32 

animal,  90,  98,  221,  230,  240,  246, 
259,  268 

electric,  133,  141,  170,  184,  187.  201, 
214,  219,  221,  264,  268 

endless-rope,  90,  247 

man,  85 

steam,  68,  82,  141 

tail-rope,  89,  259 
Hausse  formulae,  42 
Hazen  formulae,  51 
Hazleton,  Pa.,  55 
Headframes,  98 
Headings,  driving,  20-24 
Hibhing,  Minn.,  217 
Highland  Boy  mine,  Utah,  207 
Hilldale  colliery,  Pa.,  82 
Hoist-cableways,  87 
Hoisting  coal,  253,  259 

ore,  97,  98,  104,  133,  195 
Homestake  mine,  S.  Dak.,  108,  287 
Horn  Silver  mine,  Utah,  58 
Houghton  County,  Mich.,  106,  118 
Hudson  River  tunnels,  35 
Hydraulicing,  87 


T 


Illinois,  82,  238 


306 


INDEX 


Inclines.     See  Jigs  and  Slopes. 
Ingersoll-Rand  drills,  90,  96,   107,   111, 

114,  122,  184 
Ingoldsby  car,  141 
Inspiration  mine,  Ariz.,  201 
Intercoolers,  30,  32 
Iowa,  49 
Iron  Mountain,  Mich.,  192 

mine,  Mont.,  55 
Ishpenning,  Mich.,  104 
Isothermal.     See  Air. 


Jeddo-Basin  adit,  Pa.,  60 

Jeffrey  Manufacturing  Co.,  O.,  83,  253, 

278 

Jigs,  haulage,  259,  262,  263,  292 
Joplin  district,  Mo.,  26,  95 
Joveite,  12 

K 

Keystone  drill,  73,  77 
Kimberley  mine,  Africa,  224,  291 
King  mine,  Ariz.,  156 
Kinsley,  contractor,  82 
Kirby,  E.  B.,  58A 
Kirk  mine,  Mex.,  210 


Labor,  coal-mine,  229,  242,  248,  253,  265, 

266,  296 

drilling,  63,  71,  227 
ore-mine,  98,  100, 107,  111,  113,  115, 

120, 141, 174, 183,  195,  209,  216, 

218,  224,  243,  290 
surface,  69,  75,  82,  85,  296 
Ladders,  95,  488 
Lake-basins,  50,  57 
Lake  Superior,  24,  70,  214,  289 
LaSalle  County,  111.,  238 
Launders,  278,  283 
Lausanne  adit,  Pa.,  60 
Lawton,  N.  O.,  165,  288 
Lehigh  Valley  Coal  Co.,  Pa.,  43,  55,  276, 

282 

Leyner  drill,  21,  172,  173 
Locomotives.  See  Haulage. 
Longwall  system,  229,  232,  238,  242,  250, 

255,  291 

Loosening  ground,  69,  73 
Los  Pilares  mine,  Mex.,  134,  162 
Low  Moor  mines,  Va.;  207 


Ludington  mine,  Mich.,  217,  219 
Luzerne,  Pa.,  282 
Lyddite,  12 

M 

Mahanoy  City,  Pa.,  283 

Maps.  See  Surveys. 

Marion  shovel,  68,  75 

Marquette  range,  Mich.,  104,   181,   184, 

286 

Martin,  witness,  235 
Mat,  ore,  204,  290 

timber,  206,  209,  212,  214,  220,  289 
Mayari  mines,  Cuba,  62 
Menominee  Range,  Mich.,  192,  214,  217 
Mercur  mine,  Utah,  220 
Merriman  formulae,  58 
Mesabi  Range,  Minn.,  36,  64,  85,  105,  214, 

285 

Metcalf  mine,  Ariz.,  Ariz.,  128 
Mexican  mine,  Alaska,  88 
Mexico,  84,  134,  162,176,  210,  287 
Miami  mine-,  Ariz.,  165,  288 
Mid- Atlantic  States,  49 
Milling,  surface,  65,  85,  87,  88,  104,  121, 
286 

underground  104,  130,  286,  287 
Mineral  Mining  Co.,  Wis.,  192 
Mine  La  Motte,  Mo.,  91 
Minnesota  Mine,  Minn.,  121,  218 
Misfires,  16,  74 
Missouri,  26,  91,  95 
Mitchell,  M.  W.,  100,  132,  286 
Moa  mines,  Cuba,  62 
Monongahela  River  Coal  Co.,  Pa.,  272 
Montour  mines,  Pa.,  242 
Morenzi,  Ariz.,  196 
Mowry  mine,  Ariz.,  194,  289 
Multi-charging,  25,  74 


N 


Nacozari,  Mex.,  134,  162 

Negaunee,  Mich.,  181 

Nelms,  H.  J.,  264,  266,  293 

Nevada  Consolidated  mines,  Nev.,  77 

Newhouse  adit,  Colo.,  60 

Newport  mine,  Mich.,  216,  220 

Nitro-benzol,  15 

-gelatin,  11 

-glycerin,  9 
Nitryl,  8 
Nobel,  Alfred,  9,  11,  13 


INDEX 


307 


O 


Ohio  Copper  mine,  Utah,  72   191 

Oil-well.  See  Wells. 

Old  Jordan  mine,  Utah,  206 

Oliver  Mining  Co.,  Minn.,  and  Mich.,  71, 

81,  104,  181,  217 
Oneida  adit,  Pa.,  60 
Opencutting.    See    Excavator,     Milling, 

Quarrying  and  Shoveling. 
Oquirrh  Mountains,  Utah,  220 
Overhand  stoping,    28,    106,    109,    112, 

114,  116,  119,  123,  125,  129,  137, 

143,    152,    155,    157,    161,    162, 

165,  176,  202,  287 
Oversight  mine,  Mex.,  210 
Owen  County,  Ky.,  112 


Pack-walls.  See  Drywalling. 
Panel-system,  core,  100,  116,  286,  287 
longwaH,  228,  245,  250,  255,  292 
pillar,  264,  266,  269 
slicing,  206,  208,  210 
square-set,  130,  287 
Pendleton  colliery,  Eng.,  235 
Percolation.  See  Water. 
Pewabic  mine,  Mich.,  192,  289 
Philadelphia  and  Reading  Coal  Co.,  Pa., 

276 

Picrates,  12 

Pillar  and  stall.  See  Stall. 
Pillar-caving.  See  Caving. 

-drawing,    ore-body,    98,    105,    110, 
124,    156,^162,    170,    174,    178, 
204,  219,  224,  288 
seams,  261,  264,  270,  271,  273, 

275,  282,  293 
-placing,  43,  97,  114,  121,  158,  161, 

238 

-system,  seams,  228,  246,  255,  260, 
262,  264,  267,  268,  272,  281,  292, 
293 
seams, 

Pioneer  mine,  Minn.,  1&3 
Piping,  32,  61,  169,  184,  279,  283 
Pittsburg  and  Duluth  mine,  Ariz.,  132 
Pittsburg  Seam,  Pa.,  268,  272 
Plymouth  Coal  Co.,  Pa.,  276,  278 
Porter  locomotives,  32 
Portland  mine,  Colo.,  28 
Powder,  black,  7 

permissible,  8,  16 


Powder,  smokeless,  12 
Potent  ite,  13 
Preheater,  31 
Prop-placing.  See  Timbering. 

puller,  271 

Proprietary  mine,  N.  S.  W.,  149 
Providence,  Pa.,  245 
Puertocitos  mine,  Mex.,  84 
Pumping,  31,  50,  61,  219,  278 
Pyramidal  system,  116,  189 

Q. 

Quarrying,  surface,  84,  224,  286 

underground,  91,  95,  100,  286 
Quincy  mine,  Mich.,  118,  120 

R 

Rackarock,  18 

Railroad.    See    Haulage,     Cars,    Track, 

Tunnels. 
Rainfall,  15 

Rand  district,  Transvaal,  283,  293 
Ready  Bullion  mine,  Alaska,  88 
Receiver,  air,  32 
Rectang,  longwall,  239,  291 
Resistance,  line  of  least,  18 
Retreating  system,  seams,  229,  264,  268, 

272 

Rice,  C.  T.,  204 
Richardson  formula,  42 
Ricketts,  L.  D.,  199 
Rill  system,  112,  113,  115,  124,  143,  165, 

181,  287 

Rise- working,  seams,  235 
Roburite,  16 

Roof-brushing,  95,  161,  241,  242,  256 
-control,  flat,  33-37,  231,  233 

inclined,  38,  250 
-pressure,   231-237,   241,   248,   259, 

287 
-subsidence,  40,  41,  46,  121,  225,  229, 

231,  241,  269,  272,  273,  275,  282, 

284 

Room  and  Pillar.    See  Pillar. 
Room-caving.    See  Caving. 
Roosevelt  adit,  Colo.,  59 
Run-off.    See  Water. 
Russel  Foundry  Co.,  Mich.,  67 
Ruth  mine,  Nev.,  77 


Sabrero,  discoverer,  9 
Saddle-back  system,  105 


308 


INDEX 


Sampling,  71,  77,  114,  198,  207,  295 

Sampson  system,  145 

Santa  Cruz  County,  Ariz.,  194 

Scaffolds.    See  Ladders. 

Scotch  longwall,  238,  291 

Scranton,  Pa.,  276 

Scraper,  railroad,  69 

Seams,  mining.   See  Advancing,  Blasting, 
Coal,    Longwall,    Panel,    Pillar, 
Retreating,  Roof,  Slicing, 
overlying,  43,  281,  284 
recovery  of,  230,  235,  250,  272,  275, 
276,  293 

Shafts,  43,  67,  87,  88,  92,  95,  107,  184, 
195,  208,  218,  224,  230,  238 

Sharpener,  drill,  108 

Shaw  drill,  196 

Shops,  70,  238 

Shrinkage  system,  106,  111,  112,  114,  115, 
128,  155,  159,  165,  287,  288,  292 

Sibley  mine,  Minn.,  184 

Siphons,  60 

Skips,  61,  90,  98,  104,  170,  184,  216 

Slicing  system,  163,  169,  192,  196,  206, 
208,  214,  220,  224,  283,  288,  290, 
291,  292 

Slopes,  104,  246,  255 

Slush.    See  Flushing. 

Sorting,  85,  97,  103,  107,  207,  218,  276 

Soudan,  Minn.,  121,  218 

South  mine,  N.  S.  W.,  152 

South  Range  mines,  Mich.,  118 

Spoil-banks,  69,  75,  82 

Sprengel  explosives,  15 

Springing  or  squibbing.  See  Chambering. 

Spring  Valley  collieries,  111.,  238 

Square-setting,  102,  108,  110,  115,  125, 
^30,  132,  137,  145,  162,  176, 
195,  196,  207,  210,  216,  218,  219, 
287,  289,  290 

Squeeze,  40,  230,  262,  264,  275 

Squibs,  1 

Stables,  261 

Stall  and  pillar,  228,  263 

Starless  mine,  Utah,  72 

Steam-shoveling,  surface,  -64,  75,  77,  83, 

86,  285,  286 
underground,  98,  98A 

Stepped-face.   See  Longwall. 

St.  Joehead  Co.,  Mo.,  91 

St.  Louis  Prospect  Co.,  Mo.,  92 

Stoping,  24-27.  See  also  Breast,  Over- 
hand and  Underhand. 


Stripping,  67,  73,  84,  289 

Stull  system,  118,  132,  143,  217,  221 

Sub-level  system,  185,  188,  196,  200,  214, 

220,  224 

Subsidence.    See  Roof. 
Sullivan  Machine  Co.,  N.  Y.,  171,  195, 

253 

Sunderland,  Eng.,  41 
Superior  and  Boston  mine,  Ariz.,  124 
Support,  ground.     See    Roof,     Drywall- 

ing  and  Timbering. 
Surface  support,  43,  281 
Surveys,  71,  115,  265,  267 


Tamping,  3 

Tennessee  Copper  Co.,  Tenn.,  95 
Then  Shovel  Co.,  O.,  98 
Timbering,  ore-bodies,  87,  92,  95,  99, 102, 
105,  108,  115, 118,  122,  125,  132, 
143,  146,  153,  176,  186,  206,  209, 
210,  212,  214,  219,  287,  288,  290, 
292* 

seams,  227,  241,  244,  253,  269 
Tombstone,  Ariz.,  50 
Tonite,  3 

Top-slicing.    See  Slicing. 
Track,  coal,  247,  259,  262,  269,  271 

ore,  133,  163,  170,  171,  184,  195,  214 

surface,  67,  73,  80,  81,  83 
Traders'  mine,  Mich.,  87 
Transvaal,  41,  229,  283 
Trojan  powder,  81 
Tunnels,  blast,  27,  73 

railroad,  34,  35 
Two  Harbors,  Minn.,  71 

U 

Underhand  stoping,  24,  93,  95,  102,  105, 

158,  161,  192,  286,  288 
Unwatering.    See  Drainage. 
Utah  Copper  mine,  Utah,  72,  187,  289 


Ventilation,    ore-bodies,    105,    124,    179, 

183,  186,  194,  207,  209,  218,  219 

seams,  229,  230,  239,  246,  252,  257, 

261,  262,  264,  266,  267,  274,  292 

Vermillion  Range,  Minn.,  121,  183 

Veta  Grande  mine,  Mex.,  115,  210,  287 

Veteran  mine,  Nev.,  208 

Vinton  colliery,  Pa.,  250,  292 

Vipond,  Supt.,  246 


INDEX 


309 


Vulcan  shovel,  68,  75 
YV 

Walls.     See  Drywalling. 
Ward  Bros.,  sharpener,  108 
Water,  diversion,  58,  58A 

evaporation,  49 

percolation,  50 

run-off,  49 

supply,    47,  296.     See  also  Drainage 

and  Dams. 

Waugh  drill,  161,  196 
Wauviz,  J.  D.,  124 
Weigel,  Martin,  7 


Wells,  10,  51 

West  Australia,  143 

Westfillite,  16 

Westville,  N.  S.,  255,  292 

Wetter-dynamon,  16 

Whitby  mine,  Mich.,  104 

Wilkesbarre,  Pa.,  276,  282,  283 

William's  pulverizer,  279 

Winches,  247,  248 

Wolverine  mine,  Mich.,  106,  120,  287 


Zarunna  mine,  Ecuador,  124 
Zine's,  flowage  and  fracture,  46 


MINERAL  TECHNOLOGY  LIBRARY 
UNIVERSITY  OF  CALIFORNIA  LIBRARY 
BERKELEY 

Return  to  desk  from  which  borrowed. 
This  book  is  DUE  on  the  last  date  stamped  below. 


NOV  1 4 1950 
NOV  3  0 195 


,!UN  1     1952 


LD  21-100m-9,'48(B399sl6)476 


h? 


^ 


YC  33758 


